The  Metallurgy 
of  the  Common  Metals, 

Gold,  Silver,  Iron,  Copper,  Lead, 
and  Zinc 


by 


LEONARD  S.  AUSTIN 

1 1 

Professor  of  Metallurgy  and  Ore  Dressing, 
Michigan  College  of  Mines. 


Second    Edition 

Revised  and  Enlarged 


1909 

Published  by  the 
Mining  and  Scientific  Press,  San  Francisco, 

and 
The  Mining  Magazine,  London. 


COPYRIGHT,  1909 

BY 
DEWEY  PUBLISHING  COMPANY. 


ADDENDA 


Page. 

22  line  .6,  for  CO  read  CO2. 
29     "      9   for  -62'^~  read  ^A 

J>  I   T  12X07  redQ          12 

29  "  37,  for  9.77  read  9.86. 

29  "  39,  for  87%  read  77%. 

30  "  1,  for  8.5  Ib.  read  7.6  Ib. 
30  "  2,  for  0.87  read  0.85. 

30     "     4,  for  'the  table'  read  'Fig.  4';  for  'the  heat  values'  read 

'for  2000°  the  specific  heat'. 
30     "      6,  for  8.5  read  7.6 ;  for  0.251  read  0.281. 
54     Omit  the  sentence,  lines  38  and  39. 

81  line  19,  for  'in  the  manner  seen'  read  'in  the  reaction  (6) '. 
81     '•     21  and  22,  omit  reaction  (8)  and  in  its  place  put  reaction  (9). 
81     "    25,  omit  'sulphate  is  formed  thus'. 
81     "    28,  omit  'even  more';  also,  'than  the  former'. 
153  last  line,  omit  'in  the  same  general  way'. 
155  line    1,  for  chlorine  read  chlorate. 
155     "    20,  for  Fe(OH)4  read  Fe'(OH)2. 

195  Diagram,  for  5%  washes  read  0.05%  washes;  for  washes  of  25% 
and  15%  read  washes  of  0.25%  and  0.15%  ;  for  Strong 
Solution  25%  KCN  read  Strong  Solution  0.25%  KCN;  for 
Medium  Solution  15%  KCN  read  Medium  Solution  0.15%; 
KCN;  for  Weak  Solution  5%  KCN  read  Weak  Solution 
0.05%  KCN. 
198  line  14,  for  86a  read  87a. 

223  37,  for  copper  sulphide  read  copper  sulphate. 

224  "    29,  for  HgS  read  Hg2S. 

244     "      6,  f or  =  3Hg  read  =  3CuHg. 

260     "    26,  for  Na2S2O8Cu2O8  read  2Na2S203,  3Cu2S2O3. 

270     ' '      5,  for  30  to  40  read  3000  to  4000. 

270     "    18,  for  335  read  135. 

289     "      8,  for  Fe  read  4Fe. 

289     "    14,  for  - 1400  Cal.  read  +  1400  Cal. 

289  last  line,  for  +  1400  Cal.  read  +  1600  Cal. 

291  line  37,  for  1100  read  1112. 

297     ' '    10,  for  native  read  34.4%. 

306     ' '      8,  for  Fig.  130  read  Fig.  132. 

320  Insert  line  23  between  lines  5  and  6. 

323  line  30,  for  112  Ib.  read  1112  Ib. 

331  22,  for  sulphide  read  sulphate. 

332  "    17,  for  6CuSO2  read  6Cu  +  SO2. 

(OVER) 


Page. 

339  line  33,  for  Cu2S  read  Cu2O. 

353     "    next  to  last  for  'ferrous'  read  'sodium',  and  omit  'adding'. 

353  last  line,  for  'to'  read  'in'. 

354  line    1,  before  'common  salt'  insert  '200  parts  of. 

354     "      3,  after  'CuO'  insert  'and  Cu2O',  omit  'and  dissolves'. 

354  Omit  lines  5,  6,  and  7. 

354  line  10,  for  2Cu  read  Cu.     -^ 

354     "    15   Reaction    (4)    should   read   2CuCL,  +  Cu2CL  +  3Fe  - 
4Cu  +  3FeCl2. 

354     "    20,  for  one  read  three;  for  (56)'  read  (168)  ;  for  two  read 
four;  for  (126)  read  .(252). 

354     "    24,  for  3Fe2Cl0  read  2Fe.2Cl(i. 

366     "    24,  for  sulphide  read  sulphate. 

366     "    26,  for  SO2  read  2SO2. 

376     "    12,  for  157  read  159. 

380     ' '    21,  for  Pb  +  FeO  read  PbS  +  FeO. 

382     "    36,  for  C  read  CO. 

368     "       6  and  7,  for 'the  amount  of  lime  -     -  for  Baryta' read 'the 
percentage  of  magnesia  is  multiplied  by  1.4,  and  of  baryta 
by  0.4'. 

387  Charge  sheet,  last  line,  for  ^  read  ^e 

387  line    9,  for  'ratains  1%  '  read  'retains  0.85%  '. 

387     ';    10,  for  l~  or  600  Ib.  read  ^  or  680  Ib. 

387     "    13,  for  1%  of  600  Ib.  read  0.85%  of  680  Ib. 

399     "    15,  for  'they  are  frequently  contaminated  with  impurities' 
read  'blende  as  black-jack  frequently  contains  iron.' 

399     "    16,  omit  'especially  iron'. 

403     "    39,  for  0.95%  read  95%. 

403     ' '    40,  for  retained  read  recovered. 

406     "    21,  for  retort  read  condenser. 

410     "    35,  for  1500°C.  read  500°C.;  for  550°  read  85°C.;  for  boil- 
ing read  melting. 

420  "    32  and  33,  for  '  (!%<of  the  whole)   the  zinc'  read  'it'. 

421  Omit  line  2. 

440  line    2,  for  2.3  read  2.5. 

449     ' '      4,  after  ' hydrofluosilicic '  insert '  acid ',  and  for  ' is  8%  '  read 
'contains  8%'. 

469  Insert  the  sentence,  lines  18  to  22  between  lines  12  and  13. 

470  Between  lines  27  and  28  insert  'Neutral  basis,  lOc.  up  or  down'. 

471  line    1,  omit  'Neutral  basis,  lOc.  up  or  down'. 

479     "    15,  for 'operating' read 'power,  repairs,  renewals'. 

484     "    28,  for  25c.  read  25d. 

484     "    29,  for  <     X          read         X  4.86  X 


TABLE  OF  CONTENTS 


PART  I.     GENERAL. 

Section.  Page. 

1.  Ores,  Definition  and  Classification 17 

2.  Metallurgical  Treatment  of  Ores .  . . : 19 

3.  Thermo-Chemistry  as  Applied  to  Metallurgy 20 

4.  Heat  of  Formation  of  Compounds 23 

5.  Combustion    24 

6.  Temperature   of   Combustion 29 

7.  Fuels    31 

8.  By-product  Coke-Ovens  or  Retorts 40 

9.  Fuel    or   Producer-Gas 43 

10.  Refractory    Materials 47 

11.  Sampling    55 

12.  Actual  Sampling  of  Ore 56 

13.  Sampling   Metals 64 

14.  Crushing  and  Grinding 66 

PART  II.     ROASTING. 

15.  Oxidizing  Roasting 79 

16.  The  Chemistry  of  Roasting 79 

17.  Roasting  Ores  in  Lump  Form ., 84 

18.  Roasting  Ores  in  Pulverized  Condition 91 

19.  Hand-Operated  Roasters 92 

20.  Mechanical  Roasters 97 

21.  Roasting    of    Matte 110 

22.  Losses  in  Roasting Ill 

23.  Capacity  of  Furnaces  and  Cost  of  Roasting Ill 

24.  Blast  or  Pot  Roasting  of  Ores 112 

PART  III.     GOLD. 

25.  Gold    Ores 117 

26.  Stamp-Mill    Amalgamation 118 

27.  Operation  of  the  Stamp-Battery 125 

28.  General  Arrangement  of  a  Gold  Mill 130 

29.  California  and  Colorado  Practice  in  Gold-Milling 133 

30.  Cost   of    Gold-Milling 134 

31.  The   Hydrometallurgy  of  Gold 134 

32.  Chlorination  of  Gold  Ores 136 

33.  Extraction  of  the  Gold  by  Chlorination .......: 138 

34.  The  Vat  or  Plattner  Process  of  Chlorination 139 

35.  Barrel    Chlorination..  .   143 


TABLE    OF    CONTEXTS. 

Section.  Page. 

125.  Blowing-in  the  Blast-Furnace 377 

126.  Chemistry  of  the  Blast-Furnace 378 

127.  Slags  in  Silver-Lead  Smelting 381 

128.  Action  of  Various  Bases  in  Slags 382 

129.  Fuel  in  Silver-Lead  Smelting 384 

130.  Calculation  of  a  Charge ; 385 

131.  Sampling  and  Handling  Base-Bullion 390 

132.  Flue-Dust    391 

133.  Bag-House    392 

134.  Briquetting   Flue-Dust 393 

135.  Lead-Copper  Matte 395 

136.  Cost  of  Treatment  of  Lead  Ores 395 

PART  VIII.     ZINC. 

137.  Properties  of  Zinc 399 

138.  Zinc  Ores 399 

139.  Metallurgy  of  Zinc 400 

140.  Roasting  Blende 400 

141.  Chemistry  of  Roasting  Zinc  Ores 400 

142.  Roasting   Furnaces 401 

143.  The  Smelting  or  Distillation  of  Roasted  Zinc  Ores 402 

144.  The  Zinc  Furnace 4<M 

145.  Operation  of  the  Furnace 405 

146.  Manufacture  of  Retorts  and  Condensers 408 

147.  Loss   in   Distillation 409 

148.  Cost  of  Smelting 410 

149.  The  Sadtler  Process 411 

PART  IX.     REFINING. 

150.  Phenomena  Underlying  the  Refining  of  Metals 415 

1 51.  Refining  Lead  Bullion 415 

152.  Softening 416 

153.  The  Pattison  Process 419 

154.  The  Parkes  Process 420 

155.  Variations  in  Methods  of  Refining  Base-Bullion 429 

156.  Cost  of  Refining  Base-Bullion 430 

157.  Copper  Refining 430 

158.  Melting  and  Refining  Lake  Superior  Copper .   433 

159.  Electrolytic  Copper  Refining 434 

160.  Cost  of  Electrolytic   Refining 437 

161.  Refining  Impure  Spelter 438 

162.  Parting  Gold-Silver   Bars 439 

1G3.     Refining  Cast-Iron  to  make  Wrought-Iron  and  Steel 440 

164.  Puddling  Pig-iron  to  make  Wrought-Iron 441 

165.  Steel-Making  by  the  Acid  Bessemer  Process 441 

106.     The  Acid  Bessemer  Process 442 

167.  The  Basic  Open-Hearth   Process 444 

168.  The  Betts  Process  of  the  Electrolytic  Refining  of  Lead.  .  .   448 


TABLE    OF    CONTEXTS. 

PART  X.     PLANT  AND  EQUIPMENT. 

Section.  Page. 

169.  Primary  Plant  and  Equipment 453 

170.  Location   of  Works 453 

171.  Installation  of  Plant  and  Equipment 456 

172.  Handling  Materials 457 

PART  XL     COMMERCIAL. 

173.  Kinds    of    Works 467 

174.  Organization  of  a  Metallurgical  Company 467 

175.  The  Purchase  of  Ores 468 

176.  Iron    Ores*    469 

177.  Ores  Used  in  Silver-Lead  Smelting 469 

178.  Schedule  of  Copper  Ores 472 

179.  The  Operating   Department 473 

180.  Duties  at  a  Large  Gold  Stamp-Mill 478 

181.  The  Accounting  Department 478 

182.  Labor  Costs   '. 480 

183.  Profits 481 

184.  The  Selling  Department 482 

185.  Fuels  and  Metal  Market . .                                           482 


LIST  OF  ILLUSTRATIONS 


Fig.  Page. 

1.  Mahler    Bomb    Calorimeter 22 

2.  Wind    Furnace 26 

3.  Cupola  Furnace 26 

4.  Specific  Heat  of  Gases 30 

5.  Table  Showing  Genesis  of  Natural  Fuels 32 

6.  Table  of  Natural  Solid  Fuels 33 

7.  Section  of  Charcoal  Kiln ".  37 

8.  Sections  of  By-Product  and  Beehive  Coke  Ovens 39 

9.  Sectional  Elevation  of  Otto-Hoffman  By-Product  Coke-Oven  Plant..  41 

10.  Taylor  Revolving  Bottom  Gas-Producer 44 

11.  Loomis-Pettibone  Gas-Making  Plant 46 

12.  Brick  Kiln 49 

13.  Brick  Mold 52 

14.  Re-Pressing   Machine 53 

15.  Pipe    Ore-Sampler 60 

16.  Vezin  Automatic  Sampler 60 

17.  Sampling  Works   (Plan) 61 

18.  Sampling  Works    ( Elevation ) 61 

19.  Distribution  of  Silver  in  Bar  of  Base-Bullion 64 

20.  Base-Bullion  Sampling  Punch 65 

21.  Section  of  Bar  of  Ingot  Copper 65 


LIST   OF    ILLUSTRATIONS. 

Fig.  Page. 

22.  The  Blake  Ore-Crusher 70 

23.  Blake  Ore-Crusher    (Section) 71 

24.  Crushing  Rolls 72 

25.  Flow-Sheet  for  Dry  Crushing 73 

26.  Trommel  or  Cylindrical  Revolving  Screen 74 

27.  Roast- Yard  with  Trestle  (Cross  Section) 85 

28.  Roast-Yard  with  Trestle  (Longitudinal  Section) 85 

29.  Roasting  Stalls  for  Lump  Ore  (Sectional  Elevation) 88 

30.  Roasting  Stalls  for  Lump  Ore  (Plan) 89 

31.  Reverberatory  Roasting  Furnace  (Plan) 93 

32.  Reverberatory  Roasting  Furnace  with  Fuse-Box. 93 

33.  Reverberatory  Roasting  Furnace  with  Fuse-Box 95 

34.  White-Howell  Roasting  Furnace 95 

35.  Bruckner  Roasting  Furnace 98 

36.  View  of  Wethey  Roasting  Furnace 101 

37.  Cross-Section  of  Wethey  Roasting  Furnace 101 

38.  Plan  and  Elevation  of  Edwards  Roasting  Furnace 102 

39.  Cross-Section  of  Edwards  Roasting  Furnace 103 

40.  Two-Deck  Pearce-Turret  Roasting  Furnace  (Plan) 104 

41.  Two-Deck  Pearce-Turret  Roasting  Furnace  (Elevation) 105 

42.  Detail  of  Rabbles 107 

43.  McDougall  Roasting  Furnace  (Elevation) 109 

44.  Perspective  View  of  Ten-Stamp  Battery 119 

45.  Side  Elevation  of  Ten-Stamp  Battery 120 

46.  Front  Elevation  of  Ten-Stamp  Battery 121 

47.  Single-Discharge  Mortar 122 

48.  Double-Discharge    Mortar 123 

49.  Punched  Screen 124 

50.  Canda   Cam 124 

51.  Automatic    Feeder 126 

52.  Mercury  Trap 127 

53.  Clean-Up   Pan 128 

54.  Gold  Retort 130 

55.  Plan  of  Twenty-Stamp  Mill 131 

56.  Elevation  of  Twenty-Stamp  Mill 132 

57.  Table  of  Gold  Ores 137 

58.  Chlorination  Leaching  Vat 139 

59.  Chlorine  Generator 140 

60.  Section  Through  Crushing  Mill  and  Roaster 143 

61.  Chlorination    Barrel 144 

62.  Section  of  Chlorination  Barrel 145 

63.  Precipitation  Apparatus  of  Barrel  Chlorination 148 

64.  Flow-Sheet  of  Barrel  Chlorination 151 

65.  Elevation  of  Cyanide  Plant 158 

66.  Plant  of  Cyanide  Plant 159 

67.  Wooden  Leaching  Vat  Showing  False  Bottom 160 

68.  Perspective  View  of  Wooden  Leaching  Vat 161 

69.  Plan  and  Section  of  Wooden  Leaching  Vat 162 

70.  Plan  of  Steel   Leaching  Vat 163 


LIST    OF    ILLUSTRATIONS. 

Fig.  Page. 

71.  Section  of  Steel  Leaching  Vat 163 

72.  Side   Discharge   Door 164 

73.  Bottom  Discharge  Valve 165 

74.  Sections  of  Zinc  Boxes 166 

75.  Perspective  View  of  Zinc  Boxes 167 

76.  Floating  Hose  for  Gold  Solution  Tank 168 

77.  Zinc  Lathe 169 

78.  Filter  Press 171 

79.  Centrifugal   Pump 173 

80.  Elevation  of  Fifty-Ton  Cyanide  Plant 175 

81.  View  of  Fifty-Ton  Cyanide  Plant 176 

82.  Three-Compartment  Spitzkasten 177 

83.  Callow  Classifying  Cone ' 178 

84.  Double-Cone  Classifier 179 

85.  Cyanide  Plant  for  Double  Treatment. 181 

86.  Tanks   and   Butters   Distributor 182 

87a.  Blaisdell    Excavator 184 

87b.  Sand   Distributor,  Blaisdell   System 184 

88.  Butters  Filter  Tanks 186 

89.  General  Elevation  of  Butters  Filter  Plant 187 

90.  Ground  Plan  of  Butters  Filter  Plant 187 

91.  Cone  Classifiers,   Homestake  Mill 189 

92.  Elevation  of  Sand-Plant,  Homestake  Mill 190 

93.  Plan  of  Sand-Plant,  Homestake  Mill 190 

94.  Relation  Between  Mesh  and  Extraction 194 

95.  Plan  of  Cyanide  Treatment  at  El  Oro,  Mexico 195 

96.  Gates     Tube-Mill 196 

97.  Elevation  of  Tube-Mill 197 

98.  Plan  of  Operation  of  the  Maitland  Mill 200 

99.  Single-Cone  Classifiers 201 

100.  Monadnock  (Chilean)  Mill 207 

101.  Agitating  Vat 212 

102.  Wet    Silver    Mill 220 

103.  Combination    Amalgamating    Pan , 222 

104.  Eight-Foot    Settler 225 

105.  Amalgam     Safe 227 

106.  Horizontal  Retort  and  Melting  Furnace  for  Silver  Mill 228 

107.  Boss-Process    Silver   Mill    (Elevation) 230 

108.  Boss-Process  Silver  Mill    (Plan) 231 

109.  Combination   Stamp-Mill 233 

110.  Dry-Crushing  Silver  Mill   (Elevation) 239 

111.  Dry-Crushing   Silver   Mill    (Plan) 241 

112.  Flow-Sheet  of  Augustin  Process 249 

113.  Flow-Sheet  of  Ziervogel  Process 250 

114.  Plan  and  Elevation  of  Lixiviation  Plant  for  Silver  Ores 254 

115.  Sectional  Elevation  of  Plant  for  the  Lixiviation  of  Silver  Ores 255 

116.  Vat  for  Hyposulphite  Lixiviation 256 

117.  Flow-Sheet  of  Russell  Process 261 

118.  Sectional  Elevation  of  Brown  Agitator 265 


LIST    OF    ILLUSTRATIONS. 

Fig.  Page. 

119.  Sand-Treatment  Chart  and  Record 269 

120.  Iron    Blast-Furnace 275 

121.  Iron  Blast-Furnace  with  Automatic  Charging 276 

122.  Iron  Blast-Furnace,  Detailed  Section 278 

123.  Cowper  Hot-Blast  Stove   (Elevation) 281 

124.  Cowper   Hot-Blast   Stove    (Plan) 282 

125.  Blowing  Engine 285 

126.  Heyl  and  Patterson  Pig-Casting  Machine 287 

127.  Chemical   Reactions  of  Blast-Furnace 289 

128.  Copper  Smelting  Plant 301 

129.  Blast-Furnace  for  Oxidized  Copper  Ores 303 

130.  Copper  Matting   Blast-Furnace 305 

131.  View  of  Copper  Matting  Blast-Furnace 306 

132.  Steel    Water    Jackets 307 

133.  Trapped   Spout 308 

134.  Fore-Hearth     309 

135.  Slag    Pot 310 

136.  Tuyere     311 

137.  Electric  Trolley   System 326 

138.  Reverberatory   Smelting   Furnace    (Elevation) 328 

139.  Reverberatory  Smelting  Furnace    (Plan) 329 

140.  Anaconda  Reverberatory  Furnace  (Elevation) 335 

141.  Anaconda  Reverberatory  Furnace    (Plan) 335 

142.  Copper    Converter 337 

143.  Elevation  of  Smelting  and  Converting  Plant 343 

144.  Plan  of  Smelting  and  Converting  Plant 344 

145.  Wet    Pan 345 

146.  Horizontal    Blowing    Engine 346 

147.  Hydraulic    Accumulator 347 

148.  Cast-Steel    Ladle 348 

149.  Henderson-Process    Plant 357 

150.  Muffife  Roasting  Furnace 358 

151.  Ore    Bed 363 

152.  Lead-Smelting  Reverberatory  Furnace   (Plan) 365 

153.  Lead-Smelting  Reverberatory  Furnace    (Elevation) 366 

154.  American     Ore-Hearth 368 

155.  Globe  Plant 370 

156.  Silver-Lead  Blast-Furnace  (Longitudinal  Elevation) 372 

157.  Silver-Lead   Blast-Furnace    (Transverse   Elevation) 373 

158.  Perspective  View  of  Silver-Lead  Blast-Furnace 374 

159.  End  Water-Jacket  with  Knee-Bosh 375 

160.  End  Water-Jacket,  Straight 375 

161.  Cast-Iron    Side- Jacket 376 

162.  Cast-Iron  End-Jacket 376 

163.  Bag-House,  Globe  Smelting  Works 393 

164.  White    Briquetting    Press 394 

165.  Zinc-Smelting   Furnace    (Section) 404 

166.  Zinc-Smelting  Furnace    (Elevation) 404 

167.  Zinc-Smelting    Retorts 405 


LIST    OF    ILLUSTRATIONS. 

Page. 

Charge    Scoop 406 

Stamper 406 

Flow-Sheet  of  Lead  Refining 417 

Softening    Furnace 418 

Howard    Mixer 421 

Howard  Press 422 

Howard  Press  ( Section) -. . .  422 

Skimmer 423 

Molding  Market-Lead 424 

Faber  Du  Faur  Retort 425 

English  Cupelling  Furnace 426 

English  Cupelling  Furnace  (Detailed  Sections) 426 

Test  for  English  Cupelling  Furnace 428 

Stirring    Paddle 430 

Sectional  Elevation  of  Copper  Refining  Furnace 431 

Plant  of  Copper  Refining  Furnace 431 

Section  of  Bessemer  Converter  in  Upright  Position 442 

Acid  Bessemer  Blow,  American  Practice 444 

Longitudinal  Section  and  Elevation  of  Open-Hearth  Furnace 445 

Sectional  Plan  of  Open-Hearth  Furnace 446 

Chemical  Changes  in  Basic  Open-Hearth  Presses 447 

Vertical  Belt  Elevator 458 

Steel    Elevator    Bucket 459 

Single-Strand  Endless-Chain  Elevator 460 

Double-Strand   Endless-Chain   Elevator 460 

Conveying  Belt  with  Tripper 461 

Screw  Conveyor  (Quarter  Turn) 462 

Endless-Chain   Push   Conveyor..                                                                .  463 


PREFACE  TO  FIRST  EDITION 


This  outline  of  the  metallurgy  of  the  common  metals,  namely,  gold, 
silver,  iron,  copper,  lead,  and  zinc,  is  devoted  to  the  description  of 
processes  for  winning  these  metals  from  their  ores  and  then  refining 
them.  The  metallurgy  of  iron  is  treated  only  to  the  point  where 
pig-iron  is  obtained. 

Following  the  description  of  ores,  as  well  as  of  the  fuels  used 
in  smelting  them,  and  the  materials  of  which  the  furnaces  are  con- 
structed, we  come  to  sampling,  for  the  determination  of  the  exact 
value  of  the  ore  before  treatment. 

A  chapter  has  been  devoted  to  the  subject  of  thermo-chemistry  as 
applied  to  igneous  methods  of  extraction.  The  winning  or  reduction 
of  the  various  metals  is  then  taken  up  in  order,  and  is  followed  by 
a  description  of  the  methods  of  refining  them.  Attention  is  then 
given  to  commercial  considerations,  since  the  processes  must  be 
conducted  in  a  profitable  way. 

The  author  is  indebted  to  Mr.  F.  L.  Bosqui,  who  has  not  only 
read  the  manuscript,  but  has  modified  the  portion  devoted  to  the 
cyaniding  of  gold  and  silver  ores,  as  his  special  knowledge  has 
justified.  For  the  subject  matter  relating  to  the  smelting  of  silver- 
lead  and  copper  ores,  the  author  has  drawn  on  his  own  experience, 
gained  during  a  quarter  of  a  century  of  practical  work. 

L.  S.  AUSTIX. 
Houghton,  May  1,  1907. 


PREFACE  TO  SECOND  EDITION 


The  experience  gained  in  using  the  first  edition  has  suggested 
many  changes,  and  the  book  has  accordingly  been  re-written,  add- 
ing new  matter,  describing  other  processes,  and  keeping  step  with 
modern  practice. 

In  Part  I  the  subject  of  thermo-chemistry  has  been  expanded, 
and  a  table  of  heats  of  formation  given.  The  description  of  the 
cyanide  process  has  been  amplified  and  brought  up  to  date,  for 
milling  methods  are  being  rapidly  improved,  and  cyanidation  is 
having  increased  application,  especially  in  the  treatment  of  silver- 
bearing  ores.  The  metallurgy  of  zinc  has  been  treated  more  fully, 


and   particular   attention    given   to   the   principles   underlying   the 
smelting  of  zinc  ores. 

In  the  part  devoted  to  refining  there  has  been  added  the  making 
of  wrought-iron  and  steel,  the  refining  of  zinc,  and  the  electrolytic 
refining  of  lead. 

Plant  and  equipment  is  placed  in  a  separate  chapter,  while  the 
division  describing  the  economics  of  metallurgy  has  been  thrown 
into  a  more  systematic  form.  The  author  is  indebted  to  Mr.  E.  A. 
Hersam,  who  read  the  manuscript  of  the  second  edition  and  made 
numerous  suggestions  and  corrections. 

The  author  is  indebted  to  the  following  companies  for  the  use  of 
certain  of  the  illustrations  in  this  book :  Allis-Chalmers  Co.,  Mil- 
waukee, Wis. ;  Power  &  Mining  Machinery  Co.,  Cudahy,  Wis. ;  Chis- 
holm,  Matthew  &  Co.,  Colorado  Springs,  Colo. ;  F.  M.  Davis  Iron 
Works  Co.,  Denver,  Colo. ;  Stearns-Roger  Mfg.  Co.,  Denver,  Colo. : 
Pacific  Tank  Co.,  San  Francisco,  Cal. ;  Redwood  Manufacturers  Co.. 
San  Francisco,  Cal. ;  Galigher  Machinery  Co.,  Salt  Lake  City,  Utah : 
Traylor  Engineering  Co.,  Allentown,  Pa. ;  Blaisdell  Co.,  Los  Angeles; 
Cal. ;  Denver  Engineering  Works  Co.,  Denver,  Colo. ;  Trent  Engineer- 
ing &  Machinery  Co.,  Salt  Lake  City,  Utah ;  Risdon  Iron  Works,  San 
Francisco,  Cal.;  Colorado  Iron  Works  Co.,  Denver,  Colo.;  Cyanide 
Plant  Supply  Co.,  Ltd.,  London;  The  Jeffrey  Mfg.  Co.,  Columbus, 
Ohio. 

L.  S.  AUSTIX. 

Houghton,  August  1,  1909. 


PART  I.     GENERAL 


PART  I.  GENERAL. 

1.     ORES.     DEFINITION  AND  CLASSIFICATION. 

An  ore  may  be  defined  as  a  mineral  aggregate  containing  a  metal, 
or  metals,  in  sufficient  quantity  to  make  extraction  commercially 
profitable.  Minerals  or  rocks  containing  15  to  30%  iron  would  not 
be  called  iron  ore,  nor  would  we  call  a  rock  containing  2  to  3  oz. 
silver  per  ton  a  silver  ore.  Nevertheless,  the  rock  of  the  Treadwell 
mine,  on  Douglas  Island,  Alaska,  carrying  $2.50  to  $3  in  gold,  is 
called  a  gold  ore  because  it  can  be  worked  at  a  profit.  In  general, 
ores  are  named  from  their  chief  mineral  constituent,  as  lead,  copper, 
or  silver  ores,  though  they  may  contain  other  metals.  Thus  a  lead 
ore  may  contain  silver  and  gold;  a  copper  ore,  besides  copper,  may 
contain  silver,  gold,  and  even  lead.  The  appearance  of  an  ore  may 
indicate  whether  it  carries  lead,  copper,  iron,  or  zinc,  but  gold  and 
silver  are  not  always  visible,  and  the  proper  way  to  determine  their 
presence  is  by  assay.  An  ore  containing  little  lead,  say  less  than  5%, 
is  designated  a  dry  ore.  Such  ore  is  often  silicious,  but  may  possess 
value  because  it  contains  gold  and  silver.  Ores  carrying  more  than 
5  to  10%  lead  may  be  profitably  treated  for  their  lead.  Copper  ores 
containing  5  to  10%  copper  also  frequently  contain  gold  and  silver. 

Straight  or  simple  ores  contain  in  the  main  but  one. kind  of  metal, 
such  as  gold,  silver,  copper,  or  lead.  Straight  silver,  or  free-milling 
silver  ores,  are  free  from  lead  and  copper,  and  may  be  treated  by 
amalgamation.  Straight  gold  ores,  also  free-milling,  are  those  con- 
taining the  gold  in  metallic  form  and  amenable  to  amalgamation. 
Straight  or  plain  lead,  zinc,  or  copper  ores  do  not  contain  gold  or 
silver  in  quantity  sufficient  to  pay  to  separate  the  precious  metals 
from  the  base  metal.  As  an  example,  a  lead  ore  containing  4  oz. 
silver  per  ton  would  not  ordinarily  meet  the  cost  of  extracting  the 
silver.  Blister  copper  may  contain  as  much  as  16  oz.  silver  per  ton 
and  yet  not  pay  the  cost  of  electrolytic  refining  for  its  recovery. 

Mixed  ores,  or  those  containing  two  or  more  kinds  of  metal,  are 
common,  such  as  silver-gold,  silver-gold-lead,  or  lead-zinc-copper- 
silver.  When  such  ores  contain  both  copper  and  lead  it  is  puzzling 


::  THE    METALLURGY 

at  times  to  know  how  to  designate  them.  In  doubtful  cases  smelting 
companies  have  purchased  them  either  on  the  basis  of  their  lead  or 
copper  content,  under  the  plea  that,  in  extracting  one  of  these  metals, 
the  other  is  lost  or  wasted.  Lead-silver  or  lead-silver-gold  ores  are 
those  which  carry  lead  in  such  quantity  that  when  the  lead  is  re- 
covered from  them  by  smelting,  the  precious  metals  alloy  with  it, 
and  can  be  later  easily  removed  from  the  lead.  Copper-silver,  copper- 
silver-gold,  or  copper-gold  ores,  when  smelted,  yield  their  copper,  and 
this,  like  lead,  takes  up  the  precious  metals. 

Base-metal  ores. — Lead  and  copper  ores  often  contain  zinc,  anti- 
mony, arsenic,  tellurium,  or  bismuth.  These,  in  the  process  of  reduc- 
tion, alloy  with  the  principal  metal  to  its  commercial  detriment,  and 
require  expensive  after-treatment.  While  a  free-milling  ore  permits 
the  extraction  of  most  of  its  gold  or  silver  by  simple  processes  of 
grinding  and  amalgamation,  a  refractory  or  rebellious  one  requires 
preliminary  treatment  by  roasting  before  it  can  be  amalgamated; 
otherwise  it  must  be  smelted.  Even  smelting  ores  may  present  diffi- 
culties of  treatment  that  would  require  them  to  be  called  rebellious. 
A  docile  ore,  on  the  contrary,  is  one  that  may  be  easily  treated. 
Gold  and  silver  ores  containing  arsenic  or-  antimony  may  be  cited  as 
examples  of  refractory  ores. 

As  respects  treatment,  we  may  have  free-milling,  leaching,  chlori- 
dizing,  cyaniding,  or  smelting  ores.  The  latter,  or  smelting  ores, 
may  be  basic,  silicious,  dry,  coppery,  or  leady  ores ;  while  the  former 
kinds  may  be  talcose,  quartzose,  raw,  roasting,  earthy,  argillaceous, 
light,  heavy,  or  base,  all  of  which  characteristics  modify  the  mode 
of  treatment.  Among  the  iron  ores  we  may  have  bessemer  ores  or 
those  containing  a  minimum  of  phosphorus,  and  non-bessemer,  or 
those  so  high  in  phosphorus  as  to  require  treatment  in  the  basic  open- 
hearth  furnace. 

An  ore  consists  not  only  of  the  species  of  metallic  compound 
from  which  it  is  named,  but  also  of  gangue  or  waste  matter.  This 
may  often  be  its  principal  constituent,  and  may  be  earthy,  silicious, 
argillaceous,  talcose,  or  limy,  and  the  ore  may  be  composed  largely 
of  the  lighter  gangue  with  comparatively  small  quantities  of  the 
valuable  metals  scattered  or  disseminated  through  it.  When,  as  is 
often  the  case,  the  metal  is  the  heavy  part  of  the  ore,  and  the  lighter 
part  is  the  gangue,  the  ore  may  be  concentrated  or  dressed  with  a 
view  to  removing  this  gangue.  An  ore  capable  of  being  thus  treated 
is  called  a  concentrating  ore,  and  the  valuable  heavy  part  obtained 
from  it  is  called  a  'concentrate'. 

We  may  also  divide  ores  into  sulphide  and  oxidized.    As  a  mat- 


OF    THE    COMMON    METALS.  19 

ter  of  fact,  these  merge  into  one  another,  and  it  is  often  difficult  to 
decide  to  which  class  to  assign  a  given  ore.  Carbonates  are  placed 
among  the  oxidized  ores,  since,  in  smelting,  the  carbon  dioxide  is 
readily  driven  off,  leaving  the  oxide  of  the  metal. 

Grading  ore. — Miners  often  find  it  profitable  to  sort  their  ore  into 
different  grades,  such  as  shipping  or  smelting,  and  into  milling  or 
concentrating  ore,  according  to  the  after-treatment  they  propose  to 
give  it.  This  matter  is  often  an  important  one  for  the  metallurgist 
to  consider  in  deciding  upon  the  treatment  of  ore,  as,  for  example, 
in  the  case  of  a  mixed  silver  ore,  by  the  'combination  method'. 

2.     METALLURGICAL  TREATMENT  OF  ORES. 

In  winning  or  reducing  metals  from  their  ores,  the  ideal  treat- 
ment may  be  divided  into  three  stages :  Preparation,  Extraction, 
Segregation. 

(1)  Preparation. — This  has  reference  to  the  operations  by  which 
the  ore  is  fitted  to  undergo  later  treatment.     Thus,  the  ore  may  be 
broken,   crushed,   or  comminuted  to   make  the  metallic   compound 
accessible  to  a  solvent  solution.     Or  again,  the  ore  may  have  to  be 
roasted  to  render  it  capable  of  being  reduced  to  metal. 

(2)  Extraction. — This  consists  in  having  present,  or  in  adding  to 
the  prepared  ore  or  mineral,  a  collector,  the  duty  of  which  shall  be 
to  receive  or  take  up  the  metal  in  the  ore,  reducing  it  to  a  small  bulk 
from  which  the  metal  is  easily  separated  in  the  third  operation.    The 
collector  may  be  one  of  the  metals  in  molten  condition   (mercury, 
lead,  or  copper),  or  water,  or  some  solution. 

(3)  Segregation. — This  consists  in  removing,  by  precipitation  or 
concentration,  the   metal  thus  absorbed  by  the  collector,   whereby 
the  metal  itself  is  brought  into  marketable  form.     This  does  not 
necessarily  mean  refining,  for  the  metal  may  contain  impurities  and 
still  be  in  marketable  form. 

As  an  instance  of  such  a  three-fold  treatment  we  may  take  the 
extraction  of  gold  from  a  free-milling  gold  ore.  In  the  first  stage 
the  ore  is  finely  crushed  to  unlock  the  particles  of  gold.  In  the  sec- 
ond operation  these  gold  particles  are  arrested  on  a  mercury-covered 
or  amalgamated  surface,  the  mercury  acting  as  a  collector.  In  the 
final  stage,  the  collected  gold,  together  with  the  mercury  from  the 
plate,  is  recovered  by  distilling  the  mercury,  leaving  the  gold  be- 
hind. Again,  silver-bearing  sulphide  ore,  without  preliminary  roast- 
ing, may  be  melted  in  a  furnace  with  a  lead-bearing  ore.  As  the  lead 
ore  is  reduced  it  collects  the  silver  in  the  silver  ore  with  which  it  was 


20  THE    METALLURGY 

mixed,  thus  producing  a  silver-bearing  lead.  The  silver  may  be  later 
separated  from  the  lead  in  commercial  form. 

The  two  first  operations  may  be  combined  into  one.  Thus,  iron 
ore  is  treated  in  the  blast-furnace.  In  the  upper  zone,  water  and 
carbon  dioxide  are  driven  off.  The  iron  is  next  reduced,  collecting 
the  metalloids  carbon  and  silicon  to  form  pig-iron.  Finally,  these 
metalloids  may  be  separated,  and  the  pig-iron  made  into  steel. 

3.     THERMO-CHEMISTRY  AS  APPLIED  TO  METALLURGY. 

Thermo-chemistry  treats  of  the  heat  evolved  or  absorbed  as  the 
result  of  the  combustion  of  fuel,  and  the  chemical  reactions  in  metal- 
lurgy. 

In  cupellation,  air  passing  over  the  surface  of  molten  lead,  oxi- 
dizes the  lead  with  the  evolution  of  heat.  Hydrogen  and  oxygen  in 
a  closed  vessel,  ignited  by  an  electric  spark,  unite  with  much  energy, 
forming  the  vapor  of  water,  and  producing  much  heat  by  the  re- 
action. Steam  passed  through  a  glowing  coke  fire  is  decomposed, 
forming  hydrogen  and  carbon  monoxide,  but  absorbing  heat  and 
rapidly  cooling  the  fire. 

The  heat  generated  by  the  formation  of  many  chemical  com- 
pounds has  been  determined,  the  unit  by  which  it  is  measured  being 
the  heat  required  to  raise  a  unit-weight  of  water  one  degree.  The 
different  units  are  as  follows: 

One  gram  of  water  raised  1°C.,  called  the  small  calorie  (cal.). 

One  kilogram  of  water  raised  1°C.,  called  the  large  calorie  (Cal.). 

One  pound  of  water  raised  1°C.,  called  the  pound-calorie  (Ib.  cal). 

One  pound  of  water  raised  1°F.,  called  the  British  thermal  unit 
(B.  t.  u.). 

The  gram-calorie  being  too  small  for  practical  application,  it  has 
been  customary  to  use  the  large  calorie  or  Calorie,  one  thousand 
times  greater.  In  the  calculations  that  follow,  the  pound-calorie 
will  be  used,  since  the  weights  are  given  in  pounds,  and  the  numbers 
used  in  the  kilogram  system  remain  unchanged. 

The  chemical  equation 

(1)     C  +  20  =  C02 

means  that  one  equivalent  (12  Ib.)  of  carbon  is  burned  with  two 
equivalents  (32  Ib.)  of  oxygen,  producing  carbon  dioxide  (44  Ib.), 
and  evolving  97,000  pound-calories  or  8080  heat  units  per  pound  of 
carbon  burned  to  carbon  dioxide.  In  writing  equations  the  molecular 
weight  is  understood  as  in  ordinary  chemical  equations.  The  above 
may  also  be  written 


OF   THE   COMMON    METALS.  21 

(2)     C,  O2  =  97,000 

with  a  comma  to  indicate  that  the  different  molecules,  so  separated, 
unite  to  form  C02. 

If  oxygen  burns  in  presence  of  an  excess  of  highly  heated  carbon, 
then  carbon  monoxide  is  formed,  and  this  may  be  written 

(3)     C  +  O=CO 

29,000 

which  indicates  that  by  the  combination  of  the  solid  carbon  with 
the  gaseous  oxygen  29,000  calories  have  been  formed.  Likewise  this 
may  be  written 

(4)     C,  O  =  29,000 

Since  the  12  Ib.  carbon  gives  29,000  calories,  we  have,  as  the  heat 
evolved  by  the  burning  of  one  pound  of  carbon,  2440  calories.  If 
we  burn  the  CO  thus  formed  with  a  sufficient  amount  of  air,  we  have 

(5)  CO  +  0  =  C02 

68,000 

or,  as  also  written,  CO,  O  =  68,000  calories.    Were  this  written  C,  O2 
then  we  would  have  97,000  cal.     (Equation  (2)  ). 
Equation  (5)  may  again  be  written 

(6)  CO  +  O  =  C02 

29,000        97,000  =  68,000 

This  means  that  before  this  reaction  can  take  place,  the  CO  must  be 
broken  up  according  to  the  reaction, 

(7)     CO  —  C  +  O, 

29,000 
or  again 

(8)     C,  0  =  C  +  0=   -29,000, 

the  minus  sign  meaning  that  in  this  reaction  as  much  heat  has  been 
absorbed  in  the  breaking  up  as  was  earlier  evolved  in  equation  (3), 
in  which  these  elements  united. 

In  equation  (6)  the  CO  having  been  decomposed  into  C  and  O, 
they  are  forced  to  unite  with  the  O  to  form  C02,  evolving  97,000  cal. 
The  net  result  or  the  algebraic  sum  of  the  two  reactions  is  thus  68,000 
cal.  as  given  in  equation  (6). 

1.  The  amount  of  heat  needed  to  decompose  a  compound  into  its 
constituents  is  equal  to  that  evolved  when  that  compound  is  formed 
from  those  constituents.  When  a  reaction  takes  place  by  which  heat 
is  absorbed,  as  in  equation  (8),  it  is  called  '  endothermic '.  On  the 
other  hand,  when  heat  is  evolved  in  a  reaction,  as  in  equations  (1), 
(3),  and  (5),  it  is  said  to  be  'exothermic'. 


22 


THE    METALLURGY 


2.  The  heat  evolved  in  a  chemical  process  is  the  same,  whether 
it  takes  place  directly  or  in  several  steps.  Thus  in  equation  (3)  the 
carbon  is  burned  to  CO  with  the  evolution  of  29,000  calories.  The 
CO  thus  formed,  when  burned  with  additional  oxygen,  as  in  equation 
(5),  gives  68,000  calories,  and  the  sum  of  these  two  is  97,000  cal., 
the  same  as  if  the  carbon  had  been  burned  to  CO^as  per  equation  (1). 

In  comparing  reactions  (1)  and  (3),  it  may  be  said  that  in 
presence  of  an  excess  of  oxygen,  reaction  (1)  would  take  place  rather 
than  reaction  (3).  This  is  in  accordance  with  the  law  of  Berthelot, 
namely  : 

(3)    Every  reaction  which  takes  place  independently  of  the  addi- 


Fig.    1.      MAHLER  BOMB   CALORIMETER. 

tion  of  energy  from  without  the  system,  tends  to  form  the  combina- 
tion which  is  accompanied  by  the  greatest  evolution  of  heat. 

To  determine  accurately  the  heats  of  combustion  of  fuels,  or  the 
heat  of  formation  of  compounds,  the  Mahler  bomb-calorimeter,  Fig. 
1,  is  much  used.  It  consists  of  a  steel  shell  or  bomb,  marked  B,  shown 
also  on  an  enlarged  scale  at  the  right-hand  upper  corner  of  the  illus- 
tration. The  bomb,  holding  about  a  pint  and  weighing  9  lb.,  is  shown 
to  be  closed  by  a  screw-cap,  having  a  stop-cock  threaded  connection 
x  by  which  it  may  be  connected  by  a  flexible  pipe  to  a  cylinder  0, 
which  contains  compressed  oxygen  gas.  Within  the  bomb  is  sus- 
pended a  capsule  c  in  which  is  placed  a  gram  of  the  substance  to  be 
tested.  The  cap  is  then  tightly  screwed  on,  and  oxygen  gas  under 
300  lb.  per  sq.  in.  is  allowed  to  enter.  The  shell  is  placed  in  the 


OF   THE    COMMON    METALS.  23 

calorimeter  D,  which  contains  a  known  weight  of  water.  The  ther- 
mometer T  is  set  in  place,  and  the  stirrer  or  agitator  s  is  set  in  motion 
to  bring  the  whole  apparatus  to  the  same  temperature.  The  calori- 
meter D  is  placed  within  a  larger  vessel  A  covered  with  a  thick  layer 
of  felt  and  provided  with  a  thermometer  (not  shown).  The  vessel 
A  serves  also  to  support  the  bracket  G  from  which  the  stirrer  is  sus- 
pended. The  temperature  of  the  calorimeter  having  been  noted,  the 
charge  is  ignited  by  a  coil  of  the  platinum  wire  F  (See  enlarged 
view).  The  resultant  rise  in  temperature  is  noted  by  the  ther- 
mometer T.  The  total  heat  developed  with  certain  corrections,  is 
calculated  from  the  weight  of  the  calorimeter  water,  and  from  the 
rise  in  temperature.  In  those  cases  where  the  heats  of  formation  of 
oxides  or  of  silicates  are  desired,  the  result  is  accomplished  by  re- 
spectively oxidizing  or  melting  them  in  the  bomb  with  a  known 
weight  of  a  well-determined  fuel.  The  number  of  calories  evolved 
are  the  algebraic  sum  of  those  of  the  desired  reaction  and  that  of 
the  fuel. 

Following  are  the  molecular  weights  and  the  heats  of  formation 
of  some  of  the  better  known  chemical  compounds.  From  these  may 
be  estimated  the  heat  developed  in  various  reactions. 

4.  HEAT  OF  FORMATION  OF  COMPOUNDS. 

Heat  of  formation. 

Formula.  Molecular  weight.       (in  calories). 

MgO  24  +  16  =  40  143,400 

CaO  40  +  16  =  56  131,500 

A12O3  55  +  48  =  102  392,600 

MnO  55  + 16  =  71  90,900 

FeO  56  +  16  =  72  65,700 

Fe2O3  112  +  48  =  160  195,600 

PbO  207  + 16  =  223  50,800 

ZnO  65  +  16  =  81  84,800 

CuO  63.6  +  16  =  79.6  37,700 

Cu20  127.2  + 16  =  143.2  43,800 

Sb205  240  +  80  =  320  231,200 

As2O5  150  +  80  =  230  219,400 

Ag2O  216  +  16  =  232  7,000 

Na20  46  +  16  =  62  100,900 

K26  78  +  16  =  94  98,200 

Si02  28  +  32  =  60  180,000 

CO2  12  +  32  =  44  97,000 

CO  12  +  16  =  28  29,000  (gas) 


24  . 


THE    METALLURGY 


Formula. 
H20 


CaS 

ZnS 

FeS 

Cu2S 

CuS 

PbS 

SbS 


Molecular  weight. 
2  +  16  =  18 


Ba,  C,  O. 

Ca,  C,  Oa 

Mg,  C,  03 
Ca,  Si,  03 
A12,  Si,  O7 
Mn,  Si,  08 
Fe,  Si,  08 
Na2,  S,  04 
Ca,  S,  04 
Mn,  S,  O4 
Zn,  S,  04 
Fe,  S,  O4 
Fe2,  S3,  O, 
Pb,  S,  04 
H2,  S,  04 
H2,  S,  04 
Cu,  S,  04 

Ag2,;s,  o4 

x  Hg,  Au 
x  Hg,  Ag 


137 


32  =  72 
65  +  32  =  97 
56  +  32  =  88 
127.2  +  32  =  159.2 
63.6  +  32  =  95.6 
207  +  32  =  239 
240  +  96  =  336 
216  +  32  =  248 
12  +  48  =  197 
40  +  12  +  48  =  100 
24  +  12  +  48  =  84 
40  +  28  +  48  =  116 
54  +  56  +  112  =  222 

55  +  28  +  48  =  131 

56  +  28  +  48  =  132 
46  +  32  +  64  =  142 
40  +  32  +  64  =  136 
65  +  32  +  64  =  151 
65  +  32  +  64  =  161 
56  +  32  +  64  =  152 

112  +  96  +  192  =  400 
207  +  32  +  64  =  303 
2  +  32  +  64  =  98 
in  dilute  solution 
63.6  +  32  +  64  =  159.6 
216  +  32+64  =  312 
x  +  197  =  197  +  x 
x  +  108  ==  108  +  x 

5.     COMBUSTION. 


Heat  of  formation, 
(in  calories). 

70,400   (solid) 

69,000   (liquid) 

58,060  (gas) 

94,300 

43,000 

24,000 

20,300 

10,100 

20,200 

34,400 
3,000 
286,300 
273,800 
269,900 
329,350 
767,500 
276,300 
254,600 
328,100 
317,400 
249,400 
248,000 
234,900 
650,500 
215,700 
192,200 
210,200 
181,700 
167,100 
2,580 
2,470 


Combustion,  as  generally  understood,  may  be  denned  as  a  vigor- 
ous chemical  combination,  attended  with  the  production  of  light  and 
heat.  To  start  combustion,  the  fuel  must  be  (1)  brought  to  the  tem- 
perature of  ignition;  (2)  it  must  be  maintained  at  this  temperature; 
(3)  a  sufficient  supply  of  air  must  be  provided;  and  (4)  the  products 
of  combustion  must  be  removed. 

A  jet  of  gas,  burning  as  it  issues  from  a  tube,  begins  to  take  fire 


OF   THE    COMMON   METALS.  25 

one  or  more  inches  from  the  tube,  and  continues  to  burn  as  rapidly  as' 
the  molecules  of  the  gas  come  in  contact  with  those  of  the  air.  In 
an  open-hearth  furnace  a  current  of  heated  gas  and  one  of  air  mingle 
gradually,  and  do  not  become  fully  mixed  and  inflamed  until  a  few 
feet  from  the  outlet  ports.  It  is  the  same  in  a  reverberatory  furnace, 
especially  where  there  is  but  little  more  air  than  required  for  perfect 
combustion.  The  furnace,  even  if  100  ft.  long,  is  filled  with  flame 
from  end  to  end,  showing  that  gas  at  the  distant  end  is  still  com- 
bining with  air  and  burning.  If,  in  the  fire-box  of  such  a  furnace,  we 
carry  a  thick  fire,  not  less  than  18  to  24  in.  deep,  the  air  passes 
through  the  fire  freely.  The  flame  is  then  shorter,  since  the  hydro- 
carbon gas  is  speedily  consumed.  The  long  flame  is  desirable  where 
we  wish  to  extend  combustion  through  the  furnace  and  not  to  pro- 
duce so  intense  a  heat  near  the  fire.  Cases  sometimes  occur  when, 
upon  the  stopping  of  a  blast-furnace,  the  carbon  monoxide  gas  from 
the  furnace  works  back  into  the  blast-main,  and  forming  with  the 
air  an  explosive  mixture,  is  ignited  from  the  hot  fuel,  and  explodes 
with  sufficient  violence  to  break  the  main.  If  fuel-oil  be  fed  into  an 
oil-burner  or  injector,  where  it  is  broken  up  by  a  jet  of  steam  or  air 
into  a  fine  spray,  and  then  blown  into  a  red-hot  combustion  chamber, 
it  will  burn  like  gas  with  an  intense  heat.  Finely-powdered  coal 
projected  in  the  same  way,  and  with  a  sufficient  supply  of  air  (pref- 
erably pre-heated)  burns  rapidly,  resembling  gas,  and  furnishes 
abundant  heat  and  high  temperature. 

Thus,  to  promote  rapid  combustion,  the  fuel  must  be  in  such 
form  as  to  afford  plenty  of  contact  to  the  air.  A  piece  of  charcoal 
of  large  size  burns  readily,  because,  being  porous,  the  air  readily 
finds  a  way  to  penetrate  it ;  while  a  lump  of  anthracite,  being  dense, 
burns  more  slowly.  Paper  and  kindling  wood  expose  a  large  surface 
to  air,  and  hence  ignite  readily  and  burn  rapidly.  Paper  in  books, 
and  fabrics  in  bales,  burn  with  difficulty.  They  may  pass  through 
fire  only  singed  on  the  outside.  Light  lumber  and  boards  burn 
readily,  while  heavy  beams  of  wood  resist  a  severe  fire,  with  only 
superficial  charring.  A  thick  layer  of  sawdust  or  fine  coal,  thrown 
on  a  fire,  may  extinguish  it.  Hence,  in  firing  up,  such  fine  materials 
should  be  added  sparingly,  and  used  with  lump  coal  or  pieces  of 
wood,  to  make  passages  or  cavities,  through  which  the  air  may  pass. 
Finally,  as  may  be  seen,  a  common  error  made  in  fire-building  is 
that  while  an  abundant  supply  of  fuel  may  be  present,  insufficient 
provision  is  made  for  the  free  passage  of  air  through  the  fuel.  A 
good  draft  and  a  sufficiently  large  exit-flue  must  be  provided  to 
carry  away  the  products  of  combustion. 


26 


THE    METALLURGY 


Flame  is  gas  undergoing  combustion.  Soft  coal  and  wood  burn 
with  a  flame  because  the  heat  from  burning  distils,  or  drives  out,  the 
hydrocarbon  gas  which  is  formed.  Anthracite,  coke,  or  charcoal 
burns  with  little  or  no  flame ;  while  hydrogen  burns  with  a  non- 
luminous,  though  very  hot  flame. 

As  an  illustration  of  what  takes  place  in  a  deep  fire,  let  us  con- 
sider a  fire  of  glowing  coke,  Fig.  2.  Here  the  air  enters  through  the 


Fig.   2.      WIND   FURNACE. 


Fig.   3.      CUPOLA  FURNACE. 


grate-bars,  and,  at  the  first  instant,  in  contact  with  the  glowing  fuel, 
produces  carbon  dioxide, 

(1)     C  +  20  =  C02 

with  the  development  of  a  large  amount  of  heat.  Between  two  and 
four  inches  above  the  grate  (with  a  clear  fire)  we  may  expect  to 
find  the  highest  temperature.  This  forms  the  zone  1,  Fig.  2,  and 
hence  in  a  crucible  furnace  the  bottom  of  the  crucible  should  be  set 
4  in.  above  the  grate  to  get  the  full  effect  of  the  heat  of  the  fire. 

As  we  go  upward,  the  C02  in  excess  acts  on  the  glowing  carbon 
and  dissolves  or  combines  with  it  as  follows : 

(2)     C02  +  C  =  2CO 

This  reaction  is  accompanied  by  the  absorption  of  heat,  and  thus 
zone  2,  Fig.  2,  is  cooler  than  the  one  below.  In  the  zone  3,  no  reaction 
takes  place,  the  fuel  being  simply  heated  by  the  ascending  gases.  A 
little  air  may  pass  along  the  walls,  and  issuing  above  the  surface  of 


OF    THE    COMMON    METALS. 


27 


the  fuel,  and  mixing  with  the  CO  gas,  burn  a  small  portion  of  it  to 
carbon  dioxide  with  a  blue  flame  thus : 

(3)     CO  +  0  =  C02 
This  reaction  also  is  a  heat-producing  one. 

We  finally  get,  with  a  thick  fire,  a  mixture  of  gases  of  a  compo- 
sition much  like  the  following : 

N  70%,  CO  25%,  C02  2.5%,  0  0.5%,  and  H  1.0%. 
The  presence  of  the  hydrogen  is  due  to  the  decomposition  of  the 
moisture  in  the  air.  This  mixture  of  gases  can  be  made  in  a  gas-pro- 
ducer (See  Fig.  10),  and  for  that  reason  it  is  called  producer-gas. 
Because  of  its  content  of  CO  it  can  be  burned  according  to  equation 
(3),  and  used  as  a  fuel  for  any  purpose  of  heating.  To  completely 
burn  the  gas,  and  to  get  the  most  heat  from  it,  the  thickness  of  the 
fire  should  not  be  greater  than  is  shown  in  zone  1,  Fig.  2. 

Fig.  3  represents  a  foundry-cupola  charged  with  alternate  layers 
of  coke  and  pieces  of  pig-iron.  Here  the  object  is  to  melt  the  iron, 
collecting  it  in  a  pool  or  bath  at  the  bottom  in  the  crucible  of  the 
furnace,  shown  in  the  illustration.  Air  is  forced  into  the  furnace, 
and  fills  all  the  voids,  rising  through  the  charge  chiefly  in  the  pas- 
sages or  openings  offering  the  least  resistance.  In  an  iron  blast- 
furnace (See  Fig.  121)  the  rate  of  upward  velocity  approximates 
6  ft.  per  second.  As  air  meets  the  burning  coke,  combustion  takes 
place  according  to  equation  (1),  producing  a  white  heat.  This  action, 
in  a  cupola  of  36  to  48  in.  diam.,  extends  upward  about  3  ft.  from 
the  tuyeres,  the  upper  limit  of  zone  1,  Fig,  3.  As  the  gases  enter 
the  zone  2,  the  CO2  just  formed  is  decomposed  by  contact  with  the 
hot  coke  and  forms  CO,  according  to  reaction  (2),  the  change  being 
nearly  complete  at  the  upper  limit  of  zone  2.  In  the  upper  zone  no 
change  tak*>s  place  in  the  gases,  and  they  impart  their  heat  to  the 
cold  charge,  which  is  being  supplied  as  fast  as  the  ore  sinks  below 
the  required  level.  In  this  cupola,  where  the  operation  is  one  only 
of  melting,  to  attain  the  greatest  economy  of  fuel,  the  coke  should, 
be  dense,  the  pieces  large,  and  the  blast  abundant  to  supply  plenty 
of  air.  Thus,  burning  the  coke  is  deferred  to  the  last,  less  CO  is 
formed,  and  the  combustion,  performed  largely  in  zone  1,  is  more 
nearly  complete,  developing  the  largest  possible  amount  of  heat. 

From  equation  (1)  we  find,  that  to  burn  one  pound  of  carbon  to 
CO2,  and  thus  with  the  greatest  development  of  heat,  there  is  needed 
2.66  Ib.  oxygen,  or  11.6  Ib.  air,  since  air  contains  23%  oxygen  by 
weight.  At  the  sea-level  12.4  cu.  ft.  air  weigh  one  pound.  This 
makes  143.8  cu.  ft.,  or,  in  round  numbers,  150  cu.  ft.  air  per  pound 


28  THE    METALLURGY 

of  carbon.  Ordinary  coke  contains  85%  carbon,  thus  requiring  122 
cu.  ft.  air  per  pound  of  such  coke.  While  in  theory  12  Ib.  air  should 
be  sufficient  per  pound  of  coal,  it  has  been  found  that  excess  is  needed 
for  complete  combustion.  For  natural  draft,  using  a  thin  fire,  18  to 
24  Ib.  air  has  given  the  most  satisfactory  results,  and  where  air  is 
forced  into  a  closed  ash-pit,  and  through  the  fire-bed  (undergrate 
blast),  then  16  Ib.  air,  or  even  less,  is  sufficient. 

Figuring  from  equation  (2)  in  the  same  way,  we  find,  per  pound 
of  carbon  1.33  Ib.  of  oxygen  required,  or  of  air  5.79  Ib.,  equal  to  71.8 
cu.  ft.  Upon  the  basis  of  coke  containing  85%  combustible  matter, 
61  cu.  ft.  air  are  required  per  pound  of  fuel  when  burned  to  CO. 

Chimneys  or  stacks. — In  a  furnace  reaction  not  only  is  it  neces- 
sary that  the  reacting  elements  be  present  in  mutual  contact,  but 
that  the  products  of  the  reaction  be  removed  as  fast  as  formed. 
Under  conditions  other  than  this  the  reaction  ceases.  A  draft,  there- 
fore, must  be  provided,  to  carry  away  the  waste  gas,  and  to  expel 
it  into  the  atmosphere.  This  draft  may  be  natural  or  forced.  To 
insure  obtaining  a  sufficient  draft  in  a  chimney,  the  gases  must  be 
delivered  into  the  stack  while  hot.  A  temperature  of  200 °C.  is  ample 
for  this,  but  since  the  work  of  most  furnaces  is  done  at  a  tempera- 
ture higher  than  a  red-heat,  the  excess  may  be  utilized  to  generate 
steam  by  conducting  the  gases  through  waste-heat  boilers  before 
entering  the  stack. 

In  a  reverberatory-furnace,  used  for  smelting,  the  quantity  and 
intensity  of  the  heat  depend  upon  the  amount  of  coal  burned  per 
hour.  This  varies  between  18  and  40  Ib.  per  square  foot  of  grate 
area  and,  to  completely  burn  it,  there  will  be  needed  150  cu.  ft.  air 
per  pound  of  coal  consumed.  In  these  furnaces  the  gases  escape  at 
temperatures  between  300°  and  1100°C.,  and  move  with  a  velocity 
of  12  to  20  ft.  per  second.  For  illustration,  take  the  furnace  Fig.  138 
and  139,  with  a  grate-area  of  56  sq.  ft.,  a  consumption  of  30  Ib.  of 
coal  per  square  foot  of  grate-area  per  hour,  needing  252,000  cu.  ft. 
of  free  air.  Allowing  a  temperature  in  this  instance  of  1000°C.,  and 
a  draft  velocity  of  20  ft.  per  second  at  this  high  temperature,  and 
knowing  that  these  gases  expand  1/273  of  their  volume  for  each  de- 
gree above  0°C.,  we  find  the  volume  at  1000°C.  to  be  mV73273  =4.7 
times  the  volume  at  0°C.  Assuming  the  temperature  of  the  outside 
air  to  be  0°C.,  we  shall  have  as  the  volume  of  hot  gas  per  hour 
1,184,400  cu.  ft.  At  20  ft.  per  second,  or  72,000  ft.  per  hour,  this  will 
be  an  area  of  stack  of  ^  o'oo°  =  15  sq.  ft.  The  actual  area  is  16 
square  feet. 


OF    THE    COMMON    METALS.  29 

The  total  pull,  or  suction,  that  a  chimney  can  produce,  assuming 
it  to  be  filled  with  hot  gases,  is  due  simply  to  the  ascensive  force  of 
the  gas  measured  by  the  difference  between  its  weight  and  the 
weight  of  an  equal  volume  of  the  cold  air  outside.  To  maintain  the 
velocity  of  the  gas  in  the  stack,  it  has  been  found  that  a  suction,  or 
'pull',  of  0.4  to  0.8  in.  of  water,  as  measured  by  a  water-gauge,  is 
needed.  Taking  a  draft  of  0.6  in.  in  the  above  instance  and  adding 
0.1  in.  for  friction  in  the  chimney,  we  have  0.7  in.  water  equal  to 

i26xfca^3-647  lb-  Per  s(l-  ft  A  cubic  foot  air  at  °°C'  wei^hs 
0.0807  lb.  (12.4  cu.  ft.  per  lb.).  The  gas  inside  the  stack  has  a  spe- 
cific gravity  of  1.03,  weight  of  air  being  unity,  thus  making  the 
weight  when  heated  M*vu<±w  =0.0177  lb.  per  cu.  ft.  Hence  we 
have  the  difference  (0.0807—0.0177  =  0.063)  as  the  ascensive  force 
per  foot  of  height.  The  stack  should  therefore  be|^|  =  58  ft.  or 
approximately  60  ft.  high. 

From  the  above  calculation  it  appears  that  the  draft-pressure 
varies  with  the  temperature,  and  height  of  the  chimney.  The  velocity 
of  the  gas,  or  the  amount  of  air  passing  through  the  fire  per  hour 
at  a  given  temperature,  varies  as  the  square  root  of  the  height  of 
the  stack,  in  accordance  with  the  equation  V=  V2gh;  g  being  accel- 
eration due  to  gravity  and  h  head  in  feet.  Thus  a  stack  100  ft.  high 
would  increase  the  velocity  only  1.31  times  more  than  our  58-ft.  stack 
calculated  above.  The  volume  of  gas  increases  directly  with  the  tem- 
perature, while  the  velocity  varies  with  the  square  root  of  this,  hence 
there  is  a  point  of  maximum  discharge,  at  273 °C.  At  a  higher  tem- 
perature, while  the  velocity  increases,  the  weight  of  the  gas  on  the 
contrary  diminishes. 

6.     TEMPERATURE  OF  COMBUSTION. 

By  this  is  meant  the  temperature  of  gases  resulting  from  com- 
bustion under  ordinary  atmospheric  pressure.  We  can  calculate  this 
when  we  know  the  calorific  power  of  the  fuel,  and  the  total  weight 
and  mean  specific  heat  of  the  resultant  gases.  The  specific  heat  be- 
tween 0°C.  and  the  temperature  of  combustion  increases  as  is  shown 
in  the  diagram  next  shown. 

As  an  example  of  the  use  of  the  following  table,  let  us  find  the 
maximum  temperature  of  combustion  obtained  in  burning  one  pound 
of  coke  of  85%  carbon,  using  the  theoretical  amount  of  air  or  Orfr  lb. 
(since  pure  carbon  requires  11.6  lb.),  neglecting  the  loss  of  heat  in 
the  adjoining  walls  of 'the  furnace.  Since  by  weight  there  is"8£% 
nitrogen  in  air,  there  will  be,  in  the  mixed  gas  resulting  from  com- 


30 


THE    METALLURGY 


bustion,  d£  Ib.  nitrogen,  and  3.11  Ib.  C02  (0.85  Ib.  carbon,  and 
32/i2  X  0.^^=2.26  Ib.  oxygen).  We  assume  a  temperature  of  2000° 
C..  which  is  as  near  the  desired  one  as  we  can  estimate  at  this  time. 


28oo* 


o.o 


I.O 


Fig.  4.      SPECIFIC  HEAT  OF  GASES. 


From  thff'table  we  find  0*u  liuit  iciliuui  to  be,  tor  N0,  0.251  and  for 
CO2,  0.364.    Wethenha^e: 


Ib.  nitrogen   @  0.2$1  =  2.133 

3.11  Ib.  carbon-dioxide   .  .   (a)  0.364  =  1.: 


3.265 

The  number  of  calories  necessary  to  raise  the  entire  gaseous  product 
one  degree  will  then  be  3.265.  The  heat  developed  by  the  burning 
of  the  coke,  being  6800  cal.,  the  temperature  of  combustion  is  |82°6°5 
=  2080° C.  The  specific  heat  of  the  gases  at  this  temperature  being 
so  nearly  that  at. 2000° C.  we  can  use  it  in  this  calculation.  Were  the 


OF    THE    COMMON    METALS.  31 

difference  great,  we  should  have  to  use  the  specific  heat  at  the  exact 
temperature  and  calculate  again  upon  that  basis. 

To  proceed  further,  let  us  find  the  temperature  of  combustion,  or 
flame-temperature  of  carbon  monoxide  burning  to  carbon  dioxide : 
CO  +  0  =  C02  =  68,000  cal. 
28  +  16  =  44 

One  pound  of  CO  will  produce  2440  cal.,  and  will  take  0.572  Ib. 
oxygen  or -0^-  — 2.48  Ib.  air  containing  1.91  Ib.  nitrogen.  There 
is  also  1.57  Ib.  C02  produced.  Taking  the  specific  heats  from  the 
table,  Fig.  4, 

1.91  Ib.  nitrogen  @  0.287  =  0.548  cal. 

1.57  Ib.  carbon-dioxide   .  .   @  0.364  =  0.572    " 


1.120 

for  each  degree  of  rise  in  temperature.  Hence —^  =  2200 °C.  =  the 
temperature  of  combustion.  Again,  calculating  with  the  increased 
specific  heat  of  the  newly  found  temperature  (2200°),  we  find  2112 
to  be  the  exact  number  of  calories,  nearly  identical  with  that  already 
found. 

7.     FUELS. 

A  fuel  may  be  defined  as  a  solid,  liquid,  or  gaseous  substance  that 
can  be  burned  for  the  production  of  heat  for  economic  purposes. 
Fuels  can  be  divided  into  two  classes :  natural  and  artificial.  Coal 
is  a  natural  fuel;  coke  is  an  artificial  one.  The  natural  fuels  in- 
clude 'natural  gas',  the  'mineral  oils',  and  'solid  fuel'  like  wood  or 
coal.  The  solid  natural  fuels  are  believed  to  be  of  vegetable  origin. 
They  are  substances  in  some  measure  altered  from  their  original 
condition  by  heat  and  pressure,  and  range  from  wood  through  peat, 
lignite,  bituminous  or  soft  coal,  anthracite  or  hard  coal  to  graphite 
at  the  extreme.  Artificial  fuels  may  be  divided  into  the  'solid  pre- 
pared fuels'  and  'fuel  gas'.  The  solid  fuels  are  coke  and  charcoal. 

The  relation  of  the  natural  carbonaceous  substances  is  shown  in 
Fig.  5.  Here  in  a  general  way  is  illustrated  the  chemical  and  physi- 
cal changes  that  occur  in  the  formation  of  coal  from  its  organic  con- 
stituents in  plant-tissue.  These  changes  result,  finally,  under  the 
action  of  pressure  and  high  temperature,  in  graphitic  carbon,  and 
begin  by  action  upon-  wood,  leaves,  and  root-fibers  (turf  or  peat), 
passing  through  lignites  or  freshly  formed  coal,  often  brown  in 
color,  thence  to  bituminous  coal  formed  during  recorded  geological 
time,  retaining  the  volatile  constituents,  to  anthracite  where  the 
volatile  constituents  are  mostly  eliminated  by  the  heat  and  immense 


32 


THE    METALLURGY 


pressure,  and  finally  result  in  graphite  where  distillation  completes 
the  work. 

Wood. — When  freshly  cut,  wood  contains  40%  moisture  and  in 
this  condition  is  difficult  to  burn  alone ;  but  where  this  can  be  done, 
it  develops  2300  pound-calories  per  pound.  Split  into  cord-wood, 
piled,  and  dried  for  several  months,  wood  contains  20%  moisture 
and  40%  carbon.  Its  calorific  value  thereby  increases  to  3100  calo- 


ries.  Such  wood  is  classed  as  hard  when  its  specific  gravity  is  more 
than  0.55 ;  below  this  it  is  called  soft.  While  the  calorific  intensity 
of  dry  wood  is  low,  its  combustibility  is  great,  and  it  is  well  suited 
for  use  in  reverberatory  roasting-furnaces,  since  the  volatile  consti- 
tuents, rapidly  escaping,  burn  gradually,  and  make  an  extended 
flame  along  the  hearth  of  the  furnace,  heating  it  more  uniformly 
than  could  a  flameless  fuel  like  anthracite  or  coke.  In  the  outlying 
districts  of  the  western  United  States,  where  the  metallurgist  is 
dependent  on  wood  for  generating  steam,  or  roasting  ore,  the  accu- 


OF    THE    COMMON    METALS. 


33 


mulation  of  a  sufficient  supply  of  dry  wood  should  be  one  of  his  first 
cares.  In  this  his  forethought  is  well  rewarded.  He  should  pur- 
chase wood  delivered  and  corded  near  the  works ;  and  in  measuring, 
make  equitable  allowance  for  short  dimensions  or  open  piling.  Cord- 
wood  should  'cord  up'  to  10%  solid  wood. 

Peat. — This  is  the  product  of  the  slow  decay  of  plants  under 
natural  conditions  and  the  exclusion  of  air.  It  is  extremely  variable 
in  water  and  ash,  but  the  best  air-dried  peat  retains  25%  moisture, 
while  the  ash  may  vary  from  1  to  30%.  Its  calorific  power  does  not 
exceed  3000  calories  per  pound. 

Lignite. — This  fuel  occupies  an  intermediate  place  between  peat 
and  true  coal.  Four  distinct  types  can  be  specified.  These  are  :  (1) 
fossil-wood  or  fibrous  brown-coal,  which  has  a  distinctly  woody  tex- 
ture; (2)  earthy  lignite,  without  structure  and  earthy  in  fracture; 
(3)  conchoidal  lignite,  with  a  conchoidal  fracture;  and  (4)  bitum- 
inous lignite,  a  black,  shiny  fuel  having  also  a  conchoidal  fracture. 
When  recently  mined,  lignite  contains  33%  or  more  moisture,  but 
when  air-dried,  it  loses  half  or  more  of  this  (See  C  in  table  of  natural 
fuel,  Fig.  6).  The  ash  ranges  from  3  to  30%.  Lignite  burns  with 
a  long  smoky  flame.  The  calorific  power  is  variable,  being  for  C 
as  enumerated  above,  5000;  and  5700,  6500,  and  7000  calories  in 
samples  carrying  little  ash  and  so  dried  as  to  contain  but  a  low  per- 
centage of  moisture. 


Locality. 

Character   of 
Coal. 

(_i 

s 

,1 

—    03 

Ash. 

s. 

O   M 

00 

o 

.^  ^ 

'o  ^ 

15  " 

5 

* 

> 

0 

- 

\_Wood    . 

20.00 

40.00 

1.60 

3100 

B—  Peat  .  . 

24.20 

45.30 

27.  00 

3.30 

0.20 

Gallup,  N.  M  

C  —  Lignite 

12.14 

47.63 

32.81 

7.42 

Hams  Fork,  Wyo. 

D—  Lignite 

7.75 

50.60 

35.10 

6.55 

Pittsburg,  Pa  

E  —  Bituminous  .... 

3.00 

48.50 

38.25 

7.50 

2.75 

7110 

Connellsville,  Pa. 

F  —  Bituminous  

2.90 

52.  Ou 

33.60 

8.20 

1.60 

7333 

El  Moro,  Colo  

G  —  Bituminous 

0.95 

56.41 

29.82 

12.82 

0.41 

Pocahontas,  W.V. 

H  —  Bituminous  

0.69 

73.02 

19.96 

5.67 

0.66 

Bennington,  Pa... 
Colorado  

I  —  Hemi-bitum'ous 
J  —  Semi-anthracite 

1.73 
2.27 

67.03 
78  83 

23.89 
8.83 

6.69 
9.39 

0.66 

0.68 



Pennsylvania  

K—  Anthracite  

2.98 

87.13 

3.88 

5.86 

0.65 

Rhode  Island  

i.  —  Anthracite  

1.18 

85.70 

3.80 

8.52 

0.80 

M—  Graphite  

99.  (0 

1.00 

Fig.  6.   TABLE  OF  NATURAL  SOLID  FUELS. 

The  table  (Fig.  6),  gives  the  proximate  analysis  of  a  variety  of 
fuels,  the  calorific  value  of  some  of  these,  and  also  the  sulphur  con- 
tained. The  latter  element  is  of  importance  in  smelting  iron  ores, 


34  THE    METALLURGY 

since  it  must  be  eliminated  from  the  iron  which  it  would  render 
brittle. 

Bituminous  coals. — These  are  distinguished  from  lignites  by 
their  deep-black  streak,  greater  density,  and  lamellar  structure. 
They  contain  but  little  water  when  first  mined.  We  may  distinguish 
the  following  different  kinds : 

(1)  Non-coking  coal  with  a  long  flame. — These  coals  closely 
approach  lignites,  furnish  55  to  60%  pulverulent  coke,  and  give  a 
long  smoky  flame.    The  average  composition  of  the  dry  coal  is : 

Per  cent. 

Fixed    carbon    54 

Volatile  constituent   42 

Ash    4 

Water  4 

The  calorific  power  varies  from  8000  to  8500  calories. 

(2)  Coking  long-flame  gas-coal. — These  coals  give  60  to  68% 
friable,  porous  coke.    They  contain,  when  air-dried : 

Per  cent. 

Fixed    carbon    63 

Volatile  constituent   31 

Ash    6 

Water  6 

The  calorific  power  varies  from  8500  to  8800  calories. 

(3)  Furnace  coal. — These  coals  are  black,  and  never  hard.    They 
burn  with  a  long  smoky  flame,  softening  the  while,  and  swelling  in 
the  fire.     They  yield  68  to  14%  of  a  swollen  coke,  and  when  quite 
dry  are  shown  by  analysis  to  contain : 

Per  cent. 

Fixed  carbon    69.0 

Volatile   constituent    26.5 

Ash    4.5 

The  calorific  power  of  the  dry  coal  varies  from  8800  to  9300  calories. 

(4)  Coking  coal  with  a  short  flame  (semi-bituminous  coal). — 
These  yield  74  to  82%  of  a  compact  coke  and  contain,  when  dry: 

Per  cent. 

Fixed  carbon  73 

Volatile  constituent 20 

Ash    7 

The  calorific  power  varies  from  9000  to  9600  calories. 

Anthracite  coal. — (1)  Semi-anthracite. — These  coals  burn  with  a 


OF   THE    COMMON    METALS.  .35 

short  flame  and  yield  82  to  92%  of  a  pulverulent  (sometimes  fritted) 
coke.  They  are  products  of  transition  to  true  anthracite.  Their 
composition  on  a  dry  basis  may  be  stated  as : 

Per  cent. 

Fixed  carbon  83 

Volatile  constituent   11 

Ash    6 

The  calorific  power  varies  from  9200  to  9500  calories.  The  percent- 
age of  ash  ranges  from  1  sometimes  to  30,  but  seldom  exceeds  7. 

(2)  Anthracite  proper. — This  is  the  final  product  in  the  trans- 
formation of  vegetable  matter  into  coal.  It  is  of  a  jet-black  color,  of 
a  vitreous  lustre,  homogeneous  structure,  and  conchoidal  fracture. 
When  water-free  its  composition  is: 

Per  cent. 

Fixed  carbon  89 

Volatile  constituent   4 

Ash    6 

Anthracite  burns  almost  without  flame.  The  carbonaceous  residue 
after  distillation  shows  no  sign  of  coking.  In  conjunction  with  coke 
for  use  in  the  blast-furnace  for  making  pig-iron,  Pennsylvania  an- 
thracite is  used. 

Coals  of  very  different  properties  may  appear  alike  if  represented 
only  by  proximate  analyses.  The  comparative  calorific  value  may 
be  judged  of  by  Berthier's  method.  This  consists  practically  in  the 
operations  of  a  lead  assay,  using  an  excess  of  litharge,  with  a  gram 
of  the  fuel,  and  noting  the  size  of  lead  button  reduced.  One  can  also 
judge  a  good  deal  about  the  character  of  the  coal  by  coking  it  in  a 
covered  crucible  and  weighing  the  coke  produced,  judging  the  char- 
acter by  the  appearance  of  the  product.  The  proximate  analyses 
(Fig.  6),  showing  the  different  kinds  of  coal,  determine  to  which  class 
any  given  kind  belongs. 

Graphite. — This  is  of  interest,  not  as  a  fuel,  but  as  a  refractory 
material,  particularly  when  combined  with  clay.  The  analysis  given 
in  the  table  is  that  of  a  pure  form  of  graphite.  Ceylon  graphite  con- 
tains 79 A%  carbon  and  15.5%  ash  and  some  volatile  matter. 

Petroleum  or  fuel  oil. — This  is  the  most  concentrated  of  fuels, 
and,  when  the  cost  justifies,  can  be  used  not  only  for  generating 
steam,  but  for  roasting  and  melting.  It  will  be  found,  in  burning 
fuel-oil  from  various  localities,  that  the  calorific  power  is  much  the 
same  for  the  different  kinds.  Beaumont  (Texas)  oil  has  a  calorific 
power  of  10,820  calories,  and  a  specific  gravity  of  0.88  (iy3  Ib.  per 


36  THE    METALLURGY 

gallon).  Oil  can  be  burned  in  such  a  way  as  to  give,  not  only  a  high 
and  uniform  temperature,  but  also  the  oxidizing  (roasting),  or  re- 
ducing action  that  may  be  desired.  The  air  for  combustion  is  best 
pre-heated  as  well  as  the  oil,  and  it  will  be  found  advantageous  to 
inject  the  oil  under  a  high  steam-pressure.  A  mixture  of  light  and 
heavy  oils  should  not  be  used.  In  Russia  where  oil  has  been  em- 
ployed in  open-hearth  steel-furnaces  of  10  to  15  tons  capacity,  oil  to 
the  extent  of  15  to  20%  of  the  weight  of  the  charge  has  been  used. 
As  regards  comparative  costs  at  the  Selby  Smelting  &  Lead  Works, 
Vallejo  Junction,  California,  it  was  found  that  the  saving  by  using 
oil  was  40  to  60%  with  oil  at  $1.71  per  bbl.  (42  gal.),  and  coal  at  $6 
per  ton.  A  suitable  control  of  the  grade  of  the  matte  was  possible 
by  the  regulation  of  the  flame. 

Natural  gas. — In  Ohio,  Indiana,  and  Kansas,  particularly,  there 
are  districts  where  natural  gas  has  been  obtained  by  boring  for  it, 
as  for  oil.  It  is  the  most  efficient  of  natural  fuels,  having  a  calorific 
power  of  611  cal.  per  cu.  ft.,  or  27,861  cal.  per  pound.  The  following 
analysis  will  give  an  idea  of  the  composition  of  Pennsylvania  natural 
gas.  It  shows  that  natural  gas  is  composed  chiefly  of  marsh-gas  and 
hydrogen : 

Per  cent, 
by  volume. 

Carbon-dioxide    (CO2)     0.8 

Carbon-monoxide   (CO)    1.0 

Oxygen  (02)    1.1 

Ethylene    (C2H4)    1.0 

Ethane   (C2H6)    3.6 

Methane  (marsh-gas)   (CH4)   72.2 

Hydrogen    (H2) 20.7 

Charcoal. — Wood,  packed  in  a  kiln,  and  permitted  to  partly 
burn,  changes  by  distillation  of  the  volatile  portion  by  the  heat  pro- 
duced from  the  portion  burned  into  charcoal.  The  charcoal  retains 
the  form  of  the  wood  from  which  it  was  made,  but  has  a  specific 
gravity  of  only  0.2.  It  is  of  a  dull-black  color,  soils  the  fingers  but 
slightly  if  of  good  quality,  but  much  if  poor.  It  should  ring  when 
struck,  and  should  show  the  annual  rings  of  the  wood  distinctly. 
The  density  of  charcoal  varies  with  that  of  the  wood  from  which  it 
was  made,  dense  woods  giving  a  dense  charcoal.  A  heaped  bushel 
(1.5555  cu.  ft.)  weighs  14  to  16  Ib.  When  apparently  quite  dry,  char- 
coal still  contains  10%  or  more  of  moisture.  Dry  charcoal  contains 
95%  carbon,  1.5%  ash,  and  has  a  calorific  power  of  7610  pound-calo- 


OF    THE    COMMON    METALS.  37 

ries  per  pound.  Charcoal  is  used  in  iron  blast-furnaces  particularly 
in  localities  where  wood  is  abundant ;  and  it  produces  a  pure,  strong 
iron,  free  from  sulphur,  called  '  charcoal-iron '.  Charcoal  has  been 
used  also  for  silver-lead  and  copper  smelting  in  districts  difficult  of 
access.  In  these  cases  it  has  done  especially  well  when  coke  could 
be  secured  to  use  in  conjunction  with  it.  It  is,  however,  a  friable 
fuel,  making  fine  dust  sometimes  to  the  extent  of  10%  ;  and  this  'fine' 
is  apt  to  make  trouble  in  the  blast-furnace.  If  under-burned,  it  is 
heavier  and  more  dense,  and  has  a  brown  color.  Portions  of  the 
wood  found  imperfectly  burned  are  called  'brands'  and  are  returned 
for  the  next  burning. 

Charcoal  is  generally  made  in  a  kiln.  One  of  these  is  shown  in 
section  in  Fig  7,  showing  method  of  filling.  The  kiln  is  set  at  the 
foot  of  a  steep  bank  so  that  it  can  be  charged  conveniently  from 


Fig.    7.      SECTION  OF   CHARCOAL  KILN. 

above.  It  has  two  charge-doors  A  and  B.  The  first  of  the  wood  is 
conveyed  through  the  lower  door,  and  placed.  The  remainder  is 
brought  along  the  run-way  c,  and  introduced  through  the  upper  door 
B.  There  are  three  rows  of  openings,  3  by  4  in.  in  size,  spaced  2  ft. 
apart,  around  the  bottom  of  the  kiln.  The  kiln  is  lighted  at  the  lower 
door,  and  when  fairly  started,  both  openings  A  and  B  are  closed 
with  sheet-iron  doors.  These  are  tightly  luted  with  clay,  and  the  air 
is  thus  caused  to  enter  by  the  small  holes.  When  combustion  has 
progressed  sufficiently,  these  openings  are  tightly  closed,  and  the 
kiln  is  permitted  to  cool  slowly.  The  period  of  charring  or  burning 
is  eight  days,  and  the  cooling  four  days  additional.  Such  a  kiln  holds 
25  cords  of  wood  and  produces  1125  bushels  of  charcoal  weighing  16 
Ib.  per  bushel,  or  about  20%  of  the  weight  of  wood  charged. 

By-product  charcoal. — An   example   of  the   modern   method   of 


38  THE    METALLURGY 

making  by-product  charcoal  for  iron  blast-furnace  use  is  one  at  the 
Pioneer  Iron  furnace,  Marquette,  Michigan.  Here  there  are  86  kilns 
each  holding  80  cords.  The  daily  requirement  is  20  carloads,  of  16 
cords  each,  amounting  to  320  cords.  The  kiln  is  packed  full  of  wood, 
the  sheet-iron  doors  put  on  and  closed,  and  fire  is  started  at  a  man- 
hole in  the  apex  of  the  dome.  As  soon  as  combustion  gains  sufficient 
headway,  this  opening  is  closed,  and  smoke  escapes  by  way  of  a  flue 
leading  from  the  base  of  the  kiln  to  the  chimney,  continuing  thus 
until  most  of  the  aqueous  vapor  has  escaped.  At  this  stage  the 
chimney  is  closed,  and  the  vapors  pass  by  a  smoke-main  to  the  con- 
densers, the  current  being  aided  by  an  electrically  driven  fan.  The 
cold  surface  of  the  copper  tubes  of  this  condenser  precipitates  the 
condensible  portion  of  the  gas,  while  the  gas  itself  goes  on  to  the 
boilers,  where  it  is  burned  for  steam-making.  The  condensible  por- 
tion, amounting  to  41%  of  the  weight  of  the  wood,  is  called  green 
liquor  or  pyroligneous  acid,  and  consists  mostly  of  water,  but  con- 
tains also  alcohol,  tar,  ammonia  compounds,  acetone,  and  acetic  acid. 
The  tar  is  separated  in  settling  tanks,  and  the  liquor  passes  to  the 
primary  still-house.  Copper  stills  here  remove  the  vapors  of  alcohol, 
acetic  acid,  and  much  water  from  the  liquor.  The  neutralizing  tank 
receives  the  product,  and  into  this  is  mechanically  stirred  milk-of- 
lime  to  neutralize  the  acid  by  the  formation  of  acetate  of  lime.  The 
neutralized  liquor  is  allowed  to  settle,  and  the  supernatent  solution  is 
drawn  off  and  conveyed  to  the  refining-still  house.  By  fractional 
distillation  a  crude  wood  alcohol  is  obtained  here,  and  a  solution  of 
acetate  of  lime  is  left  behind  and  recovered  by  evaporating  the  solu- 
tion. The  crude  alcohol  is  then  purified  by  further  distillation  until 
a  clear  95%  wood-alcohol  is  obtained.  A  cord  of  wood  (4500  lb.), 
yields  880  lb.,  or  19.5%  dense  charcoal  of  20  lb.  per  bushel,  208  gal. 
of  pyroligneous  acid,  8  gal.  of  wood-tar,  64  lb.  gray  acetate  of  lime, 
and  4  gal.  wood  alcohol.  By  the  sale  of  the  wood-alcohol,  acetate  of 
lime  and  formaldehyde,  and  by  the  superior  quality  and  consequently 
higher  price  of  charcoal-iron,  it  has  been  possible  to  build  up  this 
industry,  where  the  supply  of  wood  is  abundant,  in  spite  of  the  seri- 
ous competition  of  iron  smelted  in  blast-furnaces  using  coke. 

Coke. — This  is  made  from  coal  in  kilns,  in  a  way  similar  to  that 
of  making  charcoal.  Bituminous  coal  which  cokes  or  fuses  at  the 
high  temperature  of  the  kiln  or  oven  is  used  for  this  purpose. 

The  raw  screenings,  in  the  example  below,  contained  much  fine 
passing  a  lV2-m.  bar-screen.  From  this,  the  residue  left  after  remov- 
ing the  lump  of  merchantable  coal,  coke  was  made.  By  washing 
the  fixed  carbon  was  increased  and  the  ash  in  the  coke  reduced  to 


OF    THE    COMMON    METALS. 


39 


14.24%.  A  part  of  the  sulphur  also  was  removed  thereby.  The  re- 
fuse was  high  in  ash,  and  low  in  fixed  carbon,  as  was  to  be  expected ; 
but  the  yield  of  washed  coal  was  85%  of  the  raw  screenings,  arid  the 
coke  70%  of  the  washed  coal.  When  the  coal  contains  slate,  'bone', 
or  pyrite,  it  is  often  improved  by  this  process  of  washing,  or  separat- 
ing the  waste-matter  by  concentrating.  An  example  of  a  semi- 
bituminous  southwestern  coal  is  shown  below : 


Raw  scr 
Washed 
Coke 

'eenings 

<! 

Moisture. 

.  .1.40 

Volatile 
;ombustible 
matter. 

19.79 
19.10 
1.39 
15.76 

Fixed 
carbon. 

60.25 
69.35 
83.47 
30.96 

Ash.     Sulphur. 

17.33         0.85 
10.24        0.52 
14.24        0.82 
50.12         0.93 

coal           .  . 

0.79 

0.43 

Refuse  or  waste   . 

.  .2.22 

A  beehive-oven,  Fig.  8  (B),  is  charged  through  a  hole  in  the  roof. 
Each  oven  holds  5  to  6  short  tons  of  coal.    The  charge  in  making  72- 


A.  B. 

Fig.  8.    SECTIONS  OF  BY-PRODUCT  (A)  AND  BEEHIVE  (B)  COKE  OVENS. 

hour  coke  is  dropped  in  the  morning  into  the  hot  oven  from  a  coal 
larry  or  car  above,  and  is  leveled  through  the  side  door,  filling  the 
oven  to  the  depth  of  26  in.  The  door  is  then  walled  up  with  dry 
brick  and  plastered  over,  but  an  opening  is  left  near  the  top,  as 
shown  in  section,  for  the  admission  of  air.  Combustion  soon  begins, 
and  a  dark  smoke  escapes  at  the  top  opening.  After  four  hours  this 
becomes  dense  and  white,  and  the  gases  ignite  or  strike,  and  flames 
issue  from  the  top.  For  twelve  hours  the  oven  burns  with  a  dull, 
smoky  flame  above  the  surface  of  the  charge.  The  flame  becomes 
bright  by  the  second  day  and  then  the  air-supply  is  partly  cut  off. 
On  the  third  day  still  less  air  is  admitted,  and  at  the  end  of  this  day 


40  THE    METALLURGY 

110  more  flames  appear  and  the  whole  interior  of  the  oven  is  red-hot. 
The  air-openings  are  now  luted,  and  the  charge  is  left  in  this  condi- 
tion until  the  morning  of  the  fourth  day  when  the  coke  is  drawn. 
The  actual  coking  is  complete  in  55  hours  and  the  whole  operation, 
from  one  charging  to  the  next,  in  72  hours.  To  draw  the  coke  the 
temporary  brick  wall  of  the  door  is  taken  down,  and  water  from  a 
hose  played  into  the  oven.  After  being  thus  cooled  on  the  surface, 
the  coke  is  pulled  out  with  a  long-handled  coke-drag  or  hook,  and 
further  cooled  with  water  while  being  withdrawn. 

The  process  of  fusing  and  coking  begins  at  the  top,  and  extends 
downward  through  the  mass  of  coal  to  the  bottom  of  the  oven,  and  the 
coke,  when  well  burned,  takes  the  form  of  primatic  masses  with  hard 
side-surfaces  of  a  silvery  steel-gray  color,  and  top  ends  soft  and 
nearly  black.  The  silvery  appearance  is  due  to  deposited  carbon, 
which  has  the  desirable  quality  of  protecting  the  coke  against  the 
action  of  the  furnace-gases.  The  black  ends  of  the  contrary  are 
readily  attacked.  A  good  coke  has  a  well-developed  cell  structure 
which  permits  the  penetration  of  the  hot  ascending  gases  in  the 
blast-furnace.  This  so  raises  the  temperature  of  the  coke  that  the 
air,  at  the  bottom  of  a  furnace  striking  it,  produces  vigorous  and 
rapid  combustion.  Other  qualities  are  purity,  uniform  quality,  and 
sufficient  coherence  for  handling.  Purity  depends  upon  a  low  ash, 
10%  being  good,  and  6  or  8  exceptionally  pure.  Coke  intended  for 
iron  blast-furnace  work,  should  not  contain  more  than  1%  sulphur 
and  commonly  less  than  0.5  to  0.8%  of  this  element.  For  lead  or 
copper  blast-furnaces  high  sulphur  does  not  greatly  matter.  'Uni- 
form quality'  means  but  a  small  amount  of  'black  ends'.  These  as 
stated,  burn  in  the  upper  part  of  the  iron  blast-furnace  by  the  ac- 
tion of  carbon-monoxide  gas.  'Coherence  in  handling'  as  is  evident, 
is  important  where  coke  must  be  transported  far,  and  re-handled  at 
the  smelting  works.  Fines  tend  to  'slow  down'  a  furnace,  but  can 
be  rejected  by  the  use  of  a  coke-fork.  The  calorific  value  of  Pitts- 
burg  coke,  containing  89%  fixed-carbon,  10%  ash,  and  1%  sulphur, 
is  7272  Ib.  cal.  per  pound. 

8.     BY-PRODUCT  COKE-OVENS  OR  RETORTS. 

The  advantage  of  the  by-product  coke-oven  is,  that  by  the  con- 
stant, high,  and  quick  heat,  it  can  coke  a  coal  that  contains  but  little 
fusible  matter.  Besides  this  the  by-products  can  be  saved  and  the 
total  yield  of  valuable  products  thereby  increased.  Two  types  of 
ovens  of  this  kind  extensively  used  in  the  United  States  are  the  Otto- 


OF   THE    COMMON   METALS. 


41 


Hoffman  and  the  Semet-Solvay.  The  general  arrangement  of  an 
Otto-Hoffman  plant  is  shown  in  Fig.  9.  Coal  is  brought  in  over  two 
tracks,  shown  at  the  left,  and  discharged  into  feed-hoppers.  It  is 
drawn  from  these  as  required,  and  conveyed-  to  two  sets  of  rolls,  one 
for  coarse,  the  other  for  fine  crushing,  and  reduced  to  a  size  of  4  to 
10-mesh.  The  crushed  coal  is  raised  by  an  inclined  elevator,  and  dis- 


42  THE    METALLURGY 

charged  into  the  main  storage  coal-bin.  This  bin  has  a  hopper-shaped 
bottom  with  several  discharge  spouts,  delivering  to  an  8-ton  larry 
which  runs  along  on  top  of  the  ovens  or  retorts,  R,R,  of  which  there 
may  be  20  to  60,  placed  side  by  side,  in  one  block  of  masonry.  Each 
retort  or  coking  chamber  is  17  in.  wide,  43  ft.  6  in.  long,  and  6  ft.  6 
in.  high,  and  is  closed  at  each  end  by  an  air-tight  cast-iron  door.  In 
Fig.  8  (A)  is  shown  a  transverse  section  of  such  a  chamber,  indi- 
cating the  interesting  lines  of  fractures  and  columnar  structure  of 
the  coke. 

Turning  again  to  the  general  view,  Fig.  9,  the  larry  is  seen  above 
the  retorts.  It  is  worked  by  an  electric  motor  and  consists  of  5  hop- 
pers supported  by  a  frame  upon  a  traveling  carriage.  From  each 
hopper  pipes  extend  downward  to  the  feed-openings  in  the  top  of  the 
chamber.  The  doors  of  the  chamber  being  closed,  and  the  chamber 
itself  hot  from  previous  operation,  a  charge  of  8  tons  of  coal  is 
dropped  in,  and  leveled  by  means  of  a  bar  inserted  through  an  open- 
ing near  the  top  of  the  door.  Distillation  at  once  begins,  and  the 
gases  are  conducted  to  condensing-chambers  to  free  them  from  cer- 
tain by-products,  such  as  tar,  ammonia,  and  benzol.  The  first  portion 
of  the  gas  is  highest  in  illuminating  power,  say  24  candle-power,  but 
later  drops  to  16  candle-power.  The  first  is,  therefore,  sent  to  the 
city-mains  for  use  as  illuminating  gas,  the  latter  reserved  to  heat  the 
chambers  by  combustion  in  flues  which  encircle  them.  The  side-walls 
are  constructed  to  provide  these  flues  for  heating  the  oven,  and  main- 
taining the  activity  of  the  distillation.  The  products  of  combustion, 
before  entering  the  stack,  go  through  a  regenerating  chamber  con- 
taining a  checker-work  of  tile,  while  air  is  pre-heated  for  combus- 
tion in  a  similar  chamber  at  the  other  side.  Thus  the  gas  is  burned 
with  highly  heated  air,  and  produces  an  intense  heat  in  the  walls 
of  the  coking-chambers.  The  reversing  valves  are  now  changed,  and 
the  currents  of  air  and  gas  caused  to  move  in  the  opposite  direction. 
The  direction  is  thus  repeatedly  alternated,  as  is  customary  in  open- 
hearth  work.  At  the  end  of  24  hours,  when  coking  is  complete,  the 
end  doors  are  opened  and  the  coke  is  pushed  out  by  means  of  a  ram 
shown  at  the  right.  The  coke  is  received  in  a  coke-quencher,  shown 
at  the  left  of  the  oven,  and  is  here  cooled  with  water  and  afterward 
discharged  into  a  car  beneath.  The  total  yield  of  coke  is  12%,  or 
6%  more,  for  the  same  coal,  than  that  of  a  beehive  oven.  The  coke 
is  hard,  dense,  and  as  reliable  as  beehive  coke  made  from  the  same 
coal,  but  has  not  the  silvery  gloss  of  the  latter. 

Costs. — The  actual  cost  of  making  coke  may  be  stated  as  50c.  per 
ton  in  the  beehive  process  and  37c.  in  by-product  ovens.  To  this 


OF    THE    COMMON    METALS.  43 

must  be  added  the  cost  of  the  1%  tons  of  coal  required.  A  beehive 
plant  operated  6  days  per  week  and  of  400-ton  daily  capacity  would 
cost  $60,000.  A  by-product  plant  of  the  same  capacity  would  cost 
$300,000.  Allowing  for  interest  and  depreciation,  the  cost  is  found  to 
be  much  the  same  for  either  process. 

9.     FUEL  OR  PRODUCER-GAS. 

Of  the  various  kinds  of  producers  used  for  making  artificial  fuel- 
gas  we  shall  consider  two,  'the  simple  producer'  and  the  'mixed-gas 
producer'. 

The  simple  producer. — These  use  ordinary  or  inferior  fuels,  such 
as  wood,  wood-refuse,  bark,  sawdust,  or  peat, 'but  generally  soft  or 
hard  coal.  We  have  shown  in  Fig.  2  and  3,  in  the  sections  of  fur- 
naces containing  fuel,  how  gas  is  produced  where  air  rises  through 
a  deep  coke  fire  and  where  fuel  is  thus  in  excess.  The  necessary  air 
may  be  supplied  by  a  natural  draft  or  by  a  fan..  The  fuel,  descend- 
ing in  the  producer,  first  is  dried  by  the  hot,  rising  gases,  then  fur- 
ther heated  until  the  volatile  matter  is  distilled,  and  finally,  as  it 
reaches  the  lowest  zone,  is  oxidized  or  burned  by  the  entering  air. 
The  residue  is  the  ash  of  the  fuel,  which  is  either  ground  out  at  the 
bottom,  as  in  the  Taylor  producer,  Fig.  10,  or  discharged  through 
the  grate  while  containing  still  some  carbon  which  is  lost.  The  es- 
caping gases  issue  at  a  temperature  of  300°  to  1000°  C.,  and  thus 
carry  away  heat. 

An  analysis  of  producer  gas  made  from  soft  coal  gave  the  follow- 
ing analysis  by  volume : 

Per  cent. 

Carbon  dioxide   (C02)    5.0 

Carbon  monoxide  (CO)    23.0 

Oxygen  (02)   0.5 

Ethylene    (C2H4)    0.5 

Methane    (CHJ 3.0 

Hydrogen    (H2)    10.0 

Nitrogen    (N2)    58.0 

Each  pound  of  coal  will  give  60  cu.  ft.  of  such  gas,  having  a  heating 
value  of  82  cal.  per  cubic  foot. 

The  mixed-gas  producer. — This  is  the  producer  commonly  used. 
In  it  a  moderate  amount  of  steam  or  water  vapor  passes  with  the  air 
into  the  burning  fuel,  and  there  re-acts  upon  the  carbon  as  follows : 
H2O  +  C  —  CO  +  H2 
58,000       29,000  =  -  29,000 


44 


THE    METALLURGY 


One  volume  of  steam  makes  one  volume  of  carbon  monoxide  and  one 
of  hydrogen.  The  water  vapor  may  be  obtained  from  the  water- 
soaked  ashes  by  evaporation  in  the  lower  part  of  the  producer,  or  as 


Fig.    10.      TAYLOR  REVOLVING  BOTTOM  GAS-PRODUCER. 


OF    THE    COMMON    METALS.  45 

in  Fig.  10,  may  be  injected  under  pressure  into  the  fire.    Steam  also 
disintegrates  the  clinker  and  facilitates  its  removal. 

(1)  The  Taylor  revolving-bottom  producer. — This  consists  of  a 
steel,  brick-lined  shell,  of  6  ft.  diam.  inside,  and  a  cast-iron,  boshed 
portion  below,  for  ashes.    The  ashes  are  supported  by  a  flat,  circular, 
iron  bottom,  which  once  in  about  six  hours  is  revolved  by  hand,  to 
grind  up  and  remove  the  ashes.    The  ashes  fall  from  the  edge  of  the 
disk  to  the  floor  of  the  ash-pit.    Up  through  the  layer  of  ashes  a  pipe 
is  seen  to  extend,  and  by  means  of  this,  both  air  and  steam  are  in- 
troduced into  the  fire.     At  the  left  is  a  steam  injector  that  is  not 
shown,  by  means  of  which  additional  air  and  steam  are  supplied  to 
the  fire,  first  by  the  horizontal  pipe,  then  by  the  vertical  one.    Peep- 
holes at  the  side  show  the  condition  of  the  fire  and  the  height  of  the 
accumulated  ashes.    The  coal  is  charged  into  a  hopper  which  is  then 
closed  air-tight.    A  slowly  revolving  vertical  shaft  provided  with  a 
distributor  at  its  lower  end,  scatters  the  coal  evenly  over  the  fire. 
The  producer-gas  leaves  by  a  large  exit  pipe  near  the  top.     Poke- 
holes  in  the  Cover  are  provided,  and  these  are  opened  occasionally  to 
stir  the  fire  when  tending  to  have  spaces  through  which  unconsumed 
air  might  pass.     The  composition  of  mixed-gas  by  volume,  made  in 
this  way  from  bituminous  coal  is  as  follows : 

Per  cent. 

Carbon  monoxide  (CO) 24.5 

Marsh-gas  (CH4)    3.6      . 

Ethylene    (C2H4)    3.2 

Carbon  dioxide  (C02)   3.7 

Hydrogen    (H2)    17.8 

Oxygen  (0,)   0.4 

Nitrogen   (N2)    (by  difference) 46.8 

(2)  The  Loomis-Pettibone  gas  apparatus. — Fig.  11  shows  a  com- 
plete plant  of  the  Loomis-Pettibone  system,  with  a  positive  gas  ex- 
'hauster,  and  intended  both  for  producer  and  water-gas.     Its  opera- 
tion is  as  follows :    Hot  fires  are  burning  in  both  producers  or  gene- 
rators, and  the  gas  exhauster  is  in  operation.    Air  is  now  drawn  up- 
ward through  generator  1,  burning  the  fuel  and  making  producer- 
gas.    This  generator  may  have  just  received  fresh  coal  at  E,  and  the 
coal-smoke,   tarry  matter,   and  producer-gas  from  it,   are  together 
drawn  down  through  the  hot  fire  in  generator  2,  being  completely 
burned  and  fixed  in  so  doing.     The  gas  now  goes  through  valve  B 
to  the  boiler  (valve  A  being  closed),  and  the  heat  is  there  absorbed. 
It  then  passes  from  the  top  of  the  boiler  through  the  pipe  shown  to 


46 


THE    METALLURGY 


the  bottom  of  the  'scrubber',  a  cylindrical  tower  of  sheet-steel,  in 
which  it  is  caused  to  pass  upward  through  pieces  of  coke  resting  upon 
perforated  trays.  The  coke  here  is  kept  wet  by  means  of  a  water- 
spray,  and  the  gas  is  thereby  cooled  and  cleaned.  Rising  to  the  top 
and  to  the  wider  part  of  the  tower,  the  gas  passes  through  a  layer 


OF    THE    COMMON"    METALS.  47 

of  fine  shavings  or  'excelsior',  to  remove  any  remaining  dust.  It 
then  is  drawn  through  the  Root  positive-blast  exhauster  W,  and 
finally  is  driven  through  pipe  Z  to  the  gasometer  for  producer-gas, 
where  it  is  stored  for  use.  The  fire  in  generator  1  having  become 
clear  and  hot,  generator  2  is  charged  afresh,  and  the  ash-pit  door 
opened.  The  gas  current  is  then  changed  from  generator  2  to  gene- 
rator 1,  through  valve  A  (valve  B  having  been  shut)  to  the  boiler, 
thence  through  the  scrubber  and  exhauster  W,  to  the  gasometer.  The 
direction  of  the  current  is  thus  changed  at  intervals.  For  making 
water-gas,  the  ash-pit  door  is  closed  and  steam  from  the  boiler  is  in- 
jected beneath  the  grate  of  the  generator  while  the  fire  is  hot.  The 
formation  of  the  water-gas  is  completed,  or  the  gas  is  'fixed'  by  caus- 
ing it  to  pass  down  through  the  other  generator,  it  having  been  found 
that  a  part  of  the  hydrogen  reverts  to  steam  without  so  doing.  The 
making  of  water-gas  cools  the  fire  and  after  a  few  minutes  the  steam 
must  be  shut  off  and  air  again  substituted.  While  water-gas  is  be- 
ing made,  it  may  go  to  the  gasometer  through  the  pipe  Z,  or,  if  de- 
sired, to  permit  it  to  go  to  the  water-gas,  holder  Z  may  be  closed  and 
Y  opened.  When  kept  separate,  water-gas  is  reserved  for  certain 
heating  operations  for  which  producer-gas,  of  lower  calorific  power, 
would  be  unsuited.  The  purge-pipe  X  is  opened  when  starting,  and 
by  this  means  air  in  the  system  is  expelled  before  gas  is  turned  into 
the  gasometer.  Steam  may  also  be  admitted  above  the  fire,  and  thus 
caused  to  pass  down  through  the  generator,  and  form  water-gas.  In 
fact,  both  air  and  steam  may  be  introduced,  either  below  or  above 
the  fires,  to  suit  the  best  conditions  of  operating. 

10.     REFRACTORY  MATERIALS. 

General. — For  the  exterior  of  furnaces  subject  to  the  action  of 
but  little  heat,  common  red  brick,  stone  laid  in  lime-mortar  or  ce- 
ment, or  concrete  is  well  suited ;  but  for  the  lining  of  furnaces  it  is 
necessary  to  use  refractory  material  to  withstand  the  high  tempera- 
ture, and  to  resist  the  scouring  and  corroding  action  of  the  molten 
contents  of  the  furnace.  At  a  temperature  below  a  red  heat  the  com- 
bined moisture  of  lime-mortar  would  be  expelled,  and  the  mortar 
in  consequence  would  crumble.  At  a  dull  red  heat  many  stones 
crack,  and  flake  off  at  the  surface  because  of  irregular  expansion. 
Sandstone,  however,  is  resistant  to  fire,  and  has  been  used  for  fur- 
nace-lining. Red  bricks,  laid  in  clay  mortar,  withstand  a  moderate 
red  heat,  but,  at  a  temperature  much  above  this,  begin  to  soften  or 
melt. 


48  THE    METALLURGY 

Refractories. — These  substances  are  infusible  at  the  high  tempera- 
tures for  which  they  are  intended.  Thus  fire-bricks  begin  only  to 
soften  at  1500°  to  1600°C.,  and  silica-brick  at  1600°  to  1700°C.  Re- 
fractories may  be  divided  into  the  three  following  classes : 

(1)  Acid  (silica-brick,  sand,  and  ganister), — These  are  used  to 
resist  the  scouring  or  corrosive  action  of  acid  slags.     Being  highly 
refractory  they  are  more  generally  used  for  roofs  or  arches  exposed 
to  the  highest  temperatures.    In  such  positions  out  of  contact  with 
the  molten  contents  of  furnaces  they  are  not  required  to  resist  a 
serious  fluxing  action. 

(2)  Neutral    (graphite,    chrome-iron,   fire-clay,    bone-ash,    and 
carbon-brick). — These  materials  well  resist  the  action   of  neutral 
slags  which  are  neither  basic  or  acid.     In  the  case  of  a  basic  open- 
hearth  furnace,  for  example,  it  is  customary  to  interpose  a  layer  of 
neutral  chrome-iron  brick  between  the  roof  of  silica-brick  and  the 
basic-lined  hearth  slightly  above  the  level  of  the  surface  of  the  mol- 
ten contents  of  the  furnace  where  it  would  be  unaffected  by  it.    Were 
silica-brick  used  in  contact  with  the  basic  lining,  they  would  re-act 
upon  the  lining  and  melt. 

(3)  Basic  (dolomite,  magnesite,  etc.) — These  are  used  where  the 
slag  or  matte  is  basic,  as  in  the  hearth  of  the  basic  open-hearth 
furnace.     Basic  slags  quickly  scour  or  corrode  an  acid,  or  even  a 
neutral  lining.    It  will  be  noticed  that  all  the  foregoing  refractories 
not  only  have  special  resistant  power  but  are  infusible.    This  is  par- 
ticularly the  case  with  carbon,  either  in  the  form  of  gas-carbon  or 
charcoal. 

We  shall  now  discuss  the  acid  refractories. 

Sand. — This  is  used  in  repairing  or  fettling  the  interior  borders 
or  walls  of  reverberatory  furnaces.  It  is  made  to  form  a  steep  bank 
extending  above  the  level  of  the  molten  bath,  to  protect  the  wall 
from  the  corrosive  action  of  the  molten  slag.  Repairs  are  made  after 
the  charge  has  been  withdrawn,  when  the  interior  sides  of  the  fur- 
nace are  exposed.  The  sand  is  thrown  in  by  means  of  shovels,  or 
placed  by  paddles  or  spoons  provided  with  16-ft.  handles,  to  deposit 
the  sand  at  the  exact  spot  required.  Sometimes  a  little  clay-ma- 
terial is  incorporated  with  the  sand  that  it  may  be  formed  into  balls. 
These  are  skilfully  thrown  across  the  furnace  through  a  door  to  an 
eroded  spot,  or  inserted  by  means  of  the  paddle,  mentioned  above, 
and  pressed  into  position  with  the  bowl  of  a  long-handled  ladle.  The 
bottoms  of  reverberatory  furnaces  are  frequently  made  of  sand  in 
layers,  and  each  layer  fired  successively,  at  the  highest  temperature 
of  the  furnace.  The  sand  thus  becomes  fritted  together,  and  hard- 


OF    THE    COMMON    METALS.  49 

ened  into  a  coherent  bed  in  this  way,  is  built  to  the  thickness  of 
perhaps  two  feet. 

Ganister. — This  is  used  for  furnace  or  converter-lining  and  is  com- 
posed of  a  mixture  of  crushed  silicious  rock  or  quartz  to  which  has 
been  added  about  15%  clayey  material  to  make  it  cohere.  For  lin- 
ing converters,  a  silicious  ore  carrying  gold  and  silver  may  be  used 
instead  of  barren  quartz  rock.  The  material  is  rapidly  eaten  or 
scoured  away  by  the  action  of  the  molten  charge,  and  the  precious 
metal  contained  enters  the  charge.  This  in  reality  results  in  a  kind 
of  ore-smelting,  performed  incidentally,  and  without  additional  cost. 

Silica  brick. — When  quartz  or  sandstone,  containing  98%  silica, 


Fig.   12.      BRICK  KILN. 

is  moistened  and  mixed  with  a  little  lime  paste  made  from  quick- 
lime, it  coheres  sufficiently  to  be  molded  into  bricks.  These  are  first 
dried  in  a  steam-heated  d:ry  ing-room,  then  carefully  placed  in  kilns 
in  open  order,  and  burned,  at  a  temperature  gradually  increasing  to 
a  white  heat.  Fig.  12  represents  a  kiln  of  the  down-draft  type.  It 
is  a  dome-shaped  oven,  15  to  25  ft.  diam.,  coal-fired  by  means  of  fire- 
places set  in  the  exterior  wall.  The  flues  within  this  wall  are  ar- 
ranged as  shown,  so  that  the  entering  flame  rises  to  the  crown  of  the 
arch,  and,  passing  downward  through  the  brick,  goes  to  the  adjoin- 
ing stack  through  flues  in  the  floor  of  the  kiln.  Thus  a  high  even 
temperature  is  obtained,  and  the  bricks  become  sufficiently  sintered 


50  THE    METALLURGY 

to  stand  handling  and  transportation,  but  are  never  as  strong  as  the 
fire-clay  brick. 

Besides  the  lime-bond  brick,  above  described,  made  by  the  addi- 
tion of  lime  to  silica,  a  clay-bond  brick,  less  refractory,  is  made  by 
the  admixture  of  four  parts  of  flint  with  one  of  clay.  This  makes  a 
stronger  brick  than  the  lime-bond.  The  composition  of  each  of  these 
kinds  of  bricks  is  as  follows : 

Lime-bond       Clay-bond 

brick.  brick. 

Per  cent.         Per  cent. 

SiO2    93.48  86.32 

A1203  3.82  11.24 

Total  fluxing  bases .     2.62  2.50 


99.92  100.06 

The  clay-bond  brick  shows  its  greater  fusibility  in  its  alumina 
and  silica  ratio,  as  will  be  seen  under  the  constitution  of  fire-bricks, 
and  the  proportion  of  alkali  is  higher  than  in  the  lime-bond  brick, 
causing  it  to  be  much  less  refractory.  Silica  bricks  withstand  the 
highest  temperatures,  and  expand  when  heated.  To  provide  for  this, 
expansion  joints  are  arranged  in  the  roof,  side  walls,  and  bridge  of 
reverberatory  furnaces,  which  close  as  the  temperature  rises.  To 
slack  off  the  tie-rods,  also,  is  another  way  to  accomplish  the  same 
purpose.  Without  this,  furnace  arches  would  bulge,  and  tie-rods 
would  break.  The  linear  expansion  of  these  bricks  when  elevated  in 
temperature  to  a  white  heat  is  2.5  per  cent. 

We  next  shall  consider  the  neutral  refractories^ 
Graphite  or  plumbago. — Pure  carbon  in  the  absence  of  air  is  per- 
manent and  infusible  at  the  highest  temperatures.  This  is  well  exem- 
plified in  the  carbon  filament  of  an  incandescent  lamp.  Even  in  the 
arc-light,  the  carbons,  though  gradually  consumed,  do  not  melt.  In 
blast-furnaces,  pulverous  carbon  accumulates,  and  forms  scaffolds, 
and  carbon-bricks,  made  of  gas-carbon,  have  been  used  with  some 
degree  of  success  for  the  bosh-lining  of  iron  blast-furnaces.  Graphite 
is  essentially  carbon,  but  contains  as  impurities  a  little  iron  and  a 
small  quantity  of  gangue  substance.  An  analysis  of  Canadian  graphite 
gives  2%  volatile  matter,  20%  ash,  and  80%  carbon.  Such  graphite 
is  used  for  graphite  or  plumbago  crucibles  and  retorts,  when  mixed 
with  45%  air-dried  clay  and  5%  sand.  Graphite,  in  these  mixtures 
is  not  only  refractory,  but  prevents  shrinking  and  cracking  when 
the  crucible  or  other  object  is  dried  after  being  formed. 


OF    THE    COMMON    METALS.  51 

Chromite  or  chrome-iron.— This  is  a  double  oxide  of  iron  and 
chromium  (FeOCr2O3)  generally  containing  a  little  gangue.  Chrome 
ore  is  made  into  bricks  by  crushing  the  ore,  mixing  with  lime  as  in 
making  silica-brick,  and  burning.  These  bricks  should  not  contain 
more  than  40%  Cr203.  Chromite  is  not  attacked  by  silicious  slags, 
and  resists  high  temperatures. 

Fire  clay,  fire  brick,  and  tile. — These  refractories  are  the  best 
known  and  the  most  used.  The  term  fire-clay  applies  to  kinds  of 
clay  capable  of  withstanding  a  high  degree  of  heat.  In  good  fire- 
clay the  total  percentage  of  fluxing  impurities,  such  as  ferric  oxide, 
lime,  magnesia,  and  the  alkalis,  is  small  (3.5%  or  less).  In  all  fire- 
clay and  water,  some  of  the  silica  is  combined  chemically  with  the 
alumina.  This  forms  a  hydrous  aluminum  silicate,  called  kaolinite. 
Further  silica  present  is  in  the  form  of  quartz  sand.  Either  kaolinite, 
or  quartz  alone,  has  a  high  fusion  point  (1850°C.),  but  in  mixture, 
the  fusion  point  is  lower,  and  this  reaches  a  minimum  at  1670°  when 
10%  kaolinite  is  present.  By  the  continued  addition  of  silica  to 
kaolinite  we  therefore  get  a  diminution  of  refractoriness  until  this 
exact  proportion  is  reached,  and  after  this,  by  continued  addition  of 
sand,  an  increase.  The  fire-clay,  accordingly,  is  most  refractory  that 
contains  the  lowest  percentage  of  fluxing  base,  and  the  least  uncom- 
bined  sand.  A  factor  further  affecting  the  refractoriness  is  the 
coarseness  of  grain.  The  New  Jersey  air-dried  clays  have  the  fol- 
lowing composition  and  refractory  qualities. 

(IV)  (V) 

Per  cent.  Per  cent. 

Kaolinite  (clay  base) 57.47  98.95 

Free  silica 40.09  0.24 

Total  fluxing  bases 2.53  0.99 


100.09  100.18 

(IV)  (V) 

Per  cent.  Per  cent. 

SiO2    67.26  45.76 

A12O3    23.36  39.05 

H20   (combined)    6.94  14.46 

Bases  (Fe203,  CaO,  alkalis) 2.53  0.99 


100.09  100.26 

Temperature  of  fusion 1670°C.         1810°C. 

The  clay  base  is  computed  as  A1203,  2SiO3  with  combined  water, 


52 


THE    METALLURGY 


The  silica  not  present  in  this  combined  form  is  regarded  as  'free'. 
It  is  seen  that  the  less  refractory  clay  (IV)  contains  more  fluxing 
base,  more  silica  and  less  alumina  than  (V)  to  account  for  its  fusi- 
bility. 

Fire-clays  are  used  not  only  for  fire-bricks  and  tiles,  but  also  for 
muffles,  retorts, 'and  clay  vessels  of  different  sorts.  The  clay  varies 
much  in  plasticity.  Clay  alone  is  unsuited  for  bricks,  since  in  burn- 
ing it  shrinks  and  cracks.  Fire-brick  manufacturers,  therefore,  em- 
ploy a  mixture  of  several  grades  of  clay,  adding  also  a  certain  per- 
centage of  coarsely  ground  fire-brick  or  quartz.  The  addition  of  this 
unshrinking  material  prevents  the  cracking  that  otherwise  would 
result.  The  assayer,  who  uses  clay  for  luting,  mixes  with  it  for  the 
same  reason  at  least  half  its  weight  of  sand.  In  the  manufacture  of 
fire-bricks  the  required  mixture  is  ground  in  a  dry-pan,  a  machine 
similar  in  construction  to  the  Chilean  mill  (See  Fig.  145),  but  pro- 


Fig.  13.     BRICK  MOLD. 

vided  with  a  grated  bottom  made  of  perforated  plates  to  discharge 
the  material  through  these  openings  when  ground  sufficiently  fine. 
Scrapers  are  placed  in  front  of  the  rollers  to  throw  material  in  their 
path,  and  the  mixture  when  ground  is  screened,,  and  further  mixed 
in  a  horizontal  plug-mill,  being  there  tempered  by  the  addition  of 
water  to  the  desired  consistence.  (See  Fig.  164.) 

The  molding  of  bricks  is  done  by  hand  or  by  machine.  If  by  hand 
the  mixture  is  brought  to  the  consistence  of  a  mud,  and  made  into 
balls  sufficiently  large  to  fill  a  six-brick  mold.  (See  Fig.  13.)  The 
mold  is  first  sanded  to  prevent  the  adhesive  mud  from  sticking.  It 
is  thrown  into  the  mold  with  force,  to  completely  fill  it,  and  the  excess 
is  cut  off  with  a  stick  or  wire,  and  the  bricks  dumped  on  a  pallet  or 
board.  The  pallets  are  placed  upon  racks,  and  air-dried  until  so 
stiff  as  to  indent  but  slightly  under  pressure  of  the  finger.  They  are 
then  put  through  a  re-pressing  machine  (Fig.  14),  where  they  are 
given  their  exact  form.  When  re-pressed,  they  are  again  placed  on 


OP   THE    COMMON    METALS.  53 

pallets  and  run  into  a  dryer  which  is  divided  into  chambers  and 
heated  by  steam  so  that  the  last  of  the  moisture  is  removed.  The 
bricks,  now  so  coherent  that  they  can  be  handled  with  little  damage, 
are  piled  in  open  order  in  the  kiln,  already  described  (See  Fig.  12), 
and  are  burned  at  a  temperature  between  1230  and  1390°C.,  requiring 
two  or  three  weeks  for  this. 

In  machine  molding,  called  the  'stiff-mud  process',  the  clay  is 
tempered  with  less  water,  and  is  much  stiffer  when  molded  than  in 
hand-molding.  The  general  form  of  the  stiff-mud  machine,  known 
as  the  auger  machine,  is  that  of  a  horizontal  cylinder,  closed  at  one 
end,  and  tapering  to  a  rectangular  outlet  at  the  other,  the  size  of  the 
brick.  Within  the  cylinder  is  a  shaft  carrying  blades  similar  to  those 


Fig.   14.     RE-PRESSING  MACHINE. 

in  a  pug-mill,  but  at  the  end  nearest  the  die,  or  outlet,  the  blades  are 
replaced  by  a  tapering  screw.  The  tempered  clay  is  fed  into  the 
cylinder  at  the  end  farthest  from  the  die.  It  is  mixed,  and  moved 
forward  by  the  blades  until  seized  by  the  screw  which  pushes  it 
through  the  die.  The  bar  of  clay  issuing  from  the  machine  is  re- 
ceived upon  a  cutting-table  and  cut  into  bricks  by  means  of  a  wire 
frame.  The  further  treatment  of  these  bricks,  with  the  drying,  re- 
pressing, and  burning,  is  like  that  of  hand-molded  bricks. 

To  resist  abrasion,  fire-bricks  must  be  hard;  to  resist  corrosion 
or  slagging,   dense ;    and  to   resist   high  temperature   and   sudden 


54  THE    METALLURGY 

changes  of  temperature,  porous  and  coarse  in  texture.  We  accord- 
ingly use  the  hard  bricks  for  door-openings,  dense  ones  for  rever- 
beratory  furnace  walls,  and  the  porous  and  coarse  ones  for  the  roofs. 
The  refractoriness  of  a  fire-brick  depends  on  the  quantity  of  the 
fluxing  bases  (especially  alkalis)  and  silica  contained,  and  on  the 
coarseness  of  the  grain. 

Bone  ash. — This  is  made  by  burning  bones  in  a  kiln  with  an  excess 
of  air,  and 'grinding  the  white  residue  to  20-mesh  size.  Organic 
matter  is  thus  removed  and  an  impure  calcium  phosphate  obtained. 
Though  a  neutral  material,  this  resists  the  action  of  litharge,  and  it 
is  accordingly  used,  not  only  in  assaying,  but  in  making  the  'tests' 
or  movable  hearths  of  the  English  cupelling-furnace  shown  in 
Fig.  178. 

Carbon  brick. — Gas  carbon,  such  as  is  used  for  arc  lights,  is  made 
into  brick  with  a  limited  amount  of  gas-tar  and  burned  in  a  kiln. 
This  brick  has  been  found  to  be  particularly  resistant  and  refractory 
in  a  reducing  atmosphere,  as  at  the  bosh  of  an  iron-furnace. 

Dolomite. — The  alkaline-earths,  lime  and  magnesia,  are  strong 
bases  and  are  resistant  to  basic  slags,  but  are  readily  fluxed  by  the 
silica  of  silicious  slags.  Quick-lime  is  infusible,  but  is  easily  affected 
by  the  moisture  of  the  air,  and  insufficiently  coherent  to  be  used  for 
making  basic  brick.  Dolomite  is  magnesian  limestone,  and  is  a  cheap 
refractory  material.  It  is  prepared  for  use  by  burning,  much  as  is 
limestone.  The  proportion  of  lime  to  magnesia  varies  in  dolomite, 
but  the  more  magnesia  the  better  for  use  as  a  refractory.  The  com- 
position of  a  typical  sample  is  as  follows  : 

Per  cent. 

CaO  31.62 

MgO    20.19 

SiO2    1.70 

FeO   1.22 

C02    45.35 


100.08 

Magnesite. — This  is  the  most  valuable  of  the  basic  materials. 
When  magnesium  carbonate  is  calcined  at  a  high  temperature  it  loses 
the  carbon  dioxide  and  the  residue  is  practically  infusible.  Mag- 
nesite is  usually  colored  dark-brown  by  the  presence  of  about  4% 


iron  oxide.    J>  nggin™nr.nf^  fryrf  Ijft1"7  rm^l  nt  n  high  temperature 
dag&aiflt4^ftbme^^  if  in  contatl  vulli  it  in  a  fmaee. 

It  is  used  both  for  basic-lined  converters  and  for  basic  open- 


OF   THE    COMMON    METALS.  55 

hearth  furnaces  where  the  slag  contains  as  little  as  15%  silica.  It 
is  used  also  as  a  lining  for  forehearths  in  copper  matting  where  it 
would  be  in  contact  with  low-grade  corrosive  matte.  The  nature  of 
the  mineral  is  shown  by  the  following  analysis : 

Per  cent. 

CaO  1.68 

MgO  42.43 

SiO2 0.92 

Fe263  and  A12O3 4.30 

CO2  and  H20 50.41 


99.74 

Other  refractory  materials. — A  mixture  of  portland  cement  2 
parts,  clay  1  part,  and  'chamotte'  or  coarsely-ground  brick  7  parts, 
moistened  and  molded  into  bricks  or  blocks,  or  used  for  patching 
furnaces,  sets  quickly  and  withstands  a  white  heat  without  disinte- 
grating. It  is  easily  made  and  especially  useful  for  rapid  repairs. 
Only  as  much  is  mixed  as  is  to  be  used  at  once. 

While  common  red  bricks  are  not  refractory,  the  least  fusible 
can  be  used  in  that  part  of  the  roof  of  a  reverberatory  furnace  where 
the  temperature  is  not  high  or  only  at  a  red  heat.  Such  bricks  are 
used  for  backing  fire-brick  structures.  As  a  general  rule  each  kind 
of  brick  should  be  laid  in  a  material  similar  to  that  of  which  it  is 
composed.  We  should  expect  slagging  to  take  place,  for  example, 
at  joints  made  of  loam-mortar  in  fire-brick.  Such  loam,  while  cheap, 
is  inferior  to  fire-clay.  An  analysis  of  good  loam  gives : 

Per  cent. 

Si02 80.99 

A1203 9.65 

Total  impurities 4.91 

Ignition  loss 4.43 


99.98 

Here  we  note  that  the  fluxing  bases  rise  to  nearly  5%,  while 
alumina  approaches  10%,  the  ratio  of  the  most  fusible  compound 
of  alumina  and  silica.  Where  the  fluxing  bases  rise  above  5%,  there 
is  risk  of  complete  melting  at  high  temperatures. 

11.     SAMPLING. 

The  principles  and  object  of  sampling. — Sampling  consists  in  ob- 
taining from  a  large  quantity  of  ore  a  small  portion  for  assay.  This 


56  THE    METALLURGY 

must  correctly  represent  the  entire  quantity  of  the  ore,  whether  it 
be  a  few  hundred  pounds  or  thousands  of  tons,  a  wagon-load,  or  a 
ship-load.  Often  we  have  a  lot  of  ore,  in  which  rich  pieces  mingle 
with  poorer  ones,  or  even  with  waste.  In  sampling  we  must  take 
this  variation  into  account  and  represent  each  part,  not  only  accord- 
ing to  its  value,  but  also  to  its  quantity.  Often  ore  is  bought  or  sold 
upon  the  results  of  sampling.  Thousands  of  dollars  are  involved  and 
cash  is  paid  for  ore  before  the  purchaser  has  treated  it.  In  other 
cases  ores  taken  by  the  reduction  works  are  treated  separately,  the 
owner  receiving  whatever  is  obtained,  a  charge  being  made  to  cover 
the  cost  of  treatment  and  the  profit  to  the  reduction  works.  In  this 
latter  case  sampling  could  be  omitted.  Similarly  at  a  mill  and  mine, 
operated  in  one  interest,  the  sampling  may  be  omitted  when  consid- 
ered an  unnecessary  expense.  Efficiency  of  the  work  is  then  de- 
termined by  the  assay  of  the  tailing. 

If  a  reduction-works  is  producing  lead,  copper,  or  zinc,  in  a  form 
ready  for  market,  the  metals  do  not  necessarily  require  to  be  sampled. 
Whenever  the  precious  metals  are  also  present  in  such  quantity  as 
to  pay  to  separate  them,  however,  the  metal  is  sampled  to  learn  the 
values  contained  before  selling  to  the  refining-works  that  is  to  effect 
the  separation.  In  blast-furnace  treatment,  ore  and  all  other  con- 
stituents of  the  charge  are  sampled,  assayed,  and  analyzed.  From 
the  data  thus  obtained,  the  charge  can  be  correctly  calculated  and 
proportioned. 

Not  only  is  it  necessary  to  ascertain  the  value  of  ores  and  of 
metals  that  result  from  metallurgical  operations,  but  as  well  the 
value  of  the  portions  rejected.  The  efficiency  of  the  work  of  the 
metallurgist  depends  upon  thorough  extraction  from  the  parts  thrown 
away.  To  be  assured  of  this,  samples  of  slag  or  tailing  are  taken  at 
frequent  intervals.  In  finding  the  value  of  a  lot  of  ore,  we  first 
weigh  the  ore,  and  base  the  assay-value  upon  the  dry  weight.  To  do 
this  we  must  determine  the  percentage  of  moisture  contained,  as 
shown  by  a  'moisture  sample'.  We  then  sample  the  ore  regularly, 
and  finally  assay  the  regular  sample.  Thus,  suppose  we  have  a  lot 
of  ore  weighing  10,800  lb.,  containing  1%  moisture  and  by  assay 
54%  lead  worth  3c.  per  pound.  Since  the  assay  is  made  on  the  dry 
weight,  we  have,  after  deducting  moisture,  10,044  lb.  ore  containing 
5424  lb.  lead  worth,  at  3c.  per  pound,  $162.72. 

12.     ACTUAL  SAMPLING  OF  ORE. 

Receiving  and  weighing. — At  reduction  works  that  purchase  ores 
(custom  works),  the  ore  arrives  either  loose,  or  in  sacks.  Whether 


OF    THE    COMMON    METALS.  57 

received  by  wagon  or  by  car,  the  vehicle  and  ore  are  weighed 
together  on  platform  scales,  thus  finding  the  'gross  weight'.  When 
the  vehicle  is  emptied,  the  weight,  called  the  'tare',  is  similarly  taken. 
The  difference  is  the  'net  weight'  or  the  'wet  weight',  and  this  is 
recorded.  When  ore  arrives  in  sacks,  the  weight  of  the  sacks  also 
is  deducted.  Often  sacked  ore  may  be  removed  to  scales  to  be 
weighed,  and  only  the  weight  of  the  sacks  deducted,  the  difference 
being  the  net  or  wet  weight.  Sacks,  if  of  sufficient  value,  are  dried 
and  returned  to  the  owner.  Railroads  often  return  them  without 
extra  charge.  Sometimes  the  sacked  ore,  if  pulverulent,  rich,  or 
frozen,  may  be  charged,  sack  and  all,  into  the  blast-furnace,  the  sack 
serving  to  retain  the  fine  contents  until  smelted,  thus  preventing  the 
loss  of  flue-dust. 

The  moisture  sample. — In  the  theory  the  moisture  sample  should 
be  taken  at  the  instant  of  weighing,  since  the  ore  may  dry  and  be- 
come lighter.  The  sample  is  taken  while  the  car  is  being  unloaded 
or  immediately  afterward.  To  represent  by  the  sample  the  ore  as 
contained  in  the  car,  holes  are  dug  at  average  points  (setting  aside 
the  dry  top  layer)  and  small  portions  are  taken  of  ore  that  appears 
to  be  of  average  moisture.  These  portions  are  put  in  a  covered  can, 
and  50  oz.  of  the  mixture  are  weighed  on  a  moisture-scale.  After 
cautiously  drying  on  a  hot-plate,  or  preferably  over  night  on  steam- 
coils,  the  50-oz.  portion  is  again  weighed,  and  the  percentage  of 
moisture  determined  by  the  loss  in  weight.  The  shipper  often  sends 
a  representative  to  witness  the  sampling  of  his  ore.  Such  a  man 
should  pay  attention  to  this  detail,  otherwise  too  high  a  percentage 
may  be  deducted  for  moisture. 

Sampling  methods  may  be  divided  into  two  class :  hand-sampling 
and  machine  or  automatic-sampling.  Any  method  of  sampling  in- 
cludes the  starting  and  finishing  operations. 

Hand-sampling. — This  includes  the  methods  called  'grab  sam- 
pling' and  'trench  sampling',  which  are  imperfect,  and  the  regular 
methods  known  as  'coning  and  quartering',  'fractional  selection', 
and  sampling  with  the  'split  shovel'. 

The  grab  sample. — This  imperfect  method  consists  in  taking  at 
uniform  distances  over  the  pile,  proportional  amounts,  broken  from 
the  lumps  and  taken  from  the  fine.  These  portions  are  mixed,  then 
reduced  in  size  by  a  method  of  hand-sampling,  later  to  be  described. 
The  imperf ectness  of  the  method  is  due  to  the  fact  that  only  the 
upper  part  of  the  pile  is  represented  in  the  sample.  The  method  is 
used  only  as  a  quick  and  inexpensive  means  of  obtaining  an  approxi- 
mate idea  of  the  value  or  composition. 


58  THE    METALLURGY 

The  trench  sample. — This  is  taken  from  ore  in  a  pile,  or  at  a  dump, 
where  an  approximate  idea  of  the  whole  is  desired.  A  trench  is  dug 
transversely  through  the  pile,  selecting  some  place  for  this  that  is 
fairly  representative  of  the  whole.  As  the  workman  advances  with 
the  cut  he  throws  the  larger  part  or  'rejected  portion'  aside,  but  re- 
serves an  aliquot  portion.  Thus  each  tenth,  twentieth,  or  hundredth 
shovelful  is  reserved  for  sample,  to  be  further  reduced  in  size  in  the 
usual  manner.  The  lower  part  of  the  pile  by  this  means  is  repre- 
sented. In  place  of  digging  a  trench,  another  method  consists  in  dig- 
ging trial  pits  at  regular  intervals  over  the  pile  or  dump  and  reserv- 
ing an  aliquot  portion  of  the  excavated  material.  The  small  reserved 
portion  of  the  whole  is  then  subjected  to  regular  sampling  methods. 
The  trench-method  and  the  trial-pit  method  are  imperfect,  since  they 
represent  only  a  part  of  the  pile. 

The  regular  or  complete  methods  of  hand-sampling,  including 
coning  and  quartering,  fractional  selection,  and  the  split-shovel 
method,  are  used  where  an  accurate  sample  that  could  be  used  with 
confidence  in  buying  or  selling  is  desired.  Of  these  methods  coning 
and  quartering  is  the  oldest. 

Coning  and  quartering. — Before  applying  this  method,  and  to  save 
unnecessary  labor,  it  is  usual  to  take  an  aliquot  portion  of  the  ore, 
say  every  tenth  shovelful,  as  the  car  or  wagon  is  being  unloaded. 
The  sample  from  a  100-ton  lot  would  then  be  10  tons.  It  is  wheeled 
to  a  crusher,  crushed  roughly  to  l^-hi-  size,  and  as  it  falls  from  the 
crusher,  wheeled  to  a  place  on  the  sampling  floor  and  dumped  in  the 
form  of  a  circle  or  ring,  leaving  a  space  of  10  ft.  diam.  inside.  The 
workmen  circle  around  this  ring,  shoveling  the  ore  to  the  apex  of  a 
cone  which  they  build  at  the  center.  Care  is  taken  that  each  shovel- 
ful is  thrown  upon  the  apex  of  this  cone.  The  ore  is  then  drawn 
down  with  shovels,  to  the  form  of  a  flat  disk  8  to  12  in.  deep.  This 
is  marked  by  diametrical  lines  into  four  equal  sections  (hence  the 
word  quartering)  of  which  two  are  left  on  the  floor  and  two  wheeled 
away.  The  reserved  sectors  are  shoveled  again  into  a  ring  and  made 
into  a  cone  thus  half  the  size  of  the  first.  This  is  again  flattened 
down,  quartered,  and  two  opposite  quarters  reserved.  The  process 
goes  on  in  this  way,  by  coning  and  quartering  the  ore,  as  long  as  any 
single  rich  piece  can  not  appreciably  raise  the  value  of  the  sample  if 
retained,  or  lower  it  if  rejected.  At  this  point  the  ore  is  crushed  to 
half-inch  size  by  means  of  rolls.  Coning  and  quartering  is  resumed 
until  it  becomes  necessary  to  crush  again,  this  time,  say,  to  wheat 
size.  The  ore  is  then  worked  down  to  a  sample  two  pounds  in  weight, 
and  this  quantity  is  ground  to  pass  through  an  80-mesh  screen.  It  is 


OF   THE    COMMON   METALS.  59 

thoroughly  mixed  by  'rolling'  on  a  sheet  of  thin  rubber  cloth;  and 
the  mixed  product  is  placed  in  4-oz.  bottles  or  manila  sample-sacks. 
These  are  sealed  and  marked  with  name  and  particulars  of  the  lot 
of  ore. 

Fractional  selection. — This  differs  from  the  quartering  method  in 
that  every  second  or  fourth  shovelful  is  reserved  for  a  sample  after 
coarse  crushing,  but  the  selected  portion  is  coned  for  the  purpose  of 
mixing.  From  the  cone  each  second  or  fourth  shovelful  is  again  re- 
served and  coned,  and  this  is  continued  until  it  becomes  necessary  to 
re-crush.  After  this,  the  reduction  proceeds  in  the  same  way  except 
that  a  smaller  shovel  is  used  toward  the  end,  to  accord  In  capacity 
with  the  size  of  the  sample. 

The  split-shovel. — By  this  method  a  shovel,  resembling  a  fork,  is 
used,  the  tines  of  which  consist  of  troughs,  each  12  in.  long  and  2  in. 
wide.  The  space  between  the  parallel  troughs  is  2  in.  wide.  Operat- 
ing on  ore  crushed  to  half -inch  diameter,  one  half  of  the  ore  shoveled 
upon  the  sampler  drops  through  the  spaces,  and  half  is  set  aside  as 
a  sample.  The  sample  is  again  divided  by  the  split-shovel,  reducing 
the  size  again  one-half,  and  so  on  until,  as  in  other  methods,  it  is 
reduced  to  the  desired  quantity. 

For  finishing  a  sample  a  riffle  sampler  is  often  used.  This  consists 
of  a  series  of  parallel  troughs  like  a  gridiron,  the  width  of  the  troughs 
being  at  least  four  times  the  diameter  of  the  largest  pieces  of  ore 
that  are  being  handled  by  means  of  it.  As  with  the  split-shovel,  one- 
half  the  ore  is  retained  in  the  troughs,  and  one-half  falls  through  and 
is  rejected.  Each  shovelful  is  evenly  scattered  upon  the  riffles,  and 
care  is  taken  not  to  heap  the  ore  above  the  troughs. 

Machine  or  automatic  sampling. — It  will  be  seen  that  the  different 
methods  of  hand-sampling,  especially  for  large  lots  of  ore  by  coning 
and  quartering,  involve  much  labor;  and  it  had  been  sought  to  over- 
come this  by  the  use  of  machinery.  A  sample,  taken  from  a  stream 
of  ore,  is  called  a  running  sample,  and  is  taken  automatically  by  a 
sampling-machine.  These  machines  are  of  two  kinds. 

(1)  Those  which  take  part  of  the  stream  all  the  time. 

(2)  Those  which  take  the  whole  stream  at  frequent  and  regular 
intervals. 

Since  the  stream  of  ore  is  not  homogeneous,  the  first  method  is 
defective  and  the  second  generally  preferred. 

Fig.  15,  called  the  pipe-sampler,  is  an  example  of  the  first  type. 
The  ore  stream  is  delivered  into  the  hopper  at  the  top,  and  is  split 
by  deflectors,  which  eject  half  the  stream  at  the  sides  of  the  tube. 


60 


THE    METALLURGY 


The  half  retained  is  caught  by  the  next  deflectors  below,  that  repeat 
the  operation,  and  the  division  continues  as  the  ore  descends.  The 
sampler  here  shown  reduces  the  quantity  of  ore  to  one-sixteenth  the 
original  amount. 


Fig.   15. 
PIPE  ORE-SAMPLER. 


Fig.    16. 
VEZIN  AUTOMATIC  SAMPLER. 


The  Vezin  sampler  (Fig.  16)  is  a  sampler  of  the  second  kind.  It 
consists  of  a  tube  carried  by  a  vertical  shaft  making  30  revolutions 
per  minute.  Attached  to  the  side  of  the  tube  and  opening  into  it  is 
a  scoop.  As  the  shaft  revolves  '  counter-clock-wise ',  it  cuts  the  stream 


OF    THE    COMMON    METALS. 


61 


of  ore  that  is  falling  from  the  inclined  feed-chute.  The  ore^  thus 
intercepted,  falls  through  the  tube  and  becomes  the  sample,  while  the 
rejected  portion,  falling  in  the  main  hopper,  is  delivered  by  a  pipe  to 


Fig.   17.      SAMPLING  WORKS   (PLAN). 


Fig.   18.      SAMPLING  WORKS  (ELEVATION). 


62 


THE    METALLURGY 


the  bin.  In  the  plan  of  a  sampling  works  (Fig.  17)  three  automatic 
samplers  are  shown,  the  two  first  of  which  have  double  scoops.  The 
ore  is  crushed  finer  after  passing  each  sampler,  and  by  the  time  three 
successive  cuts  have  been  made  the  sample  is  reduced  to  one  two- 
hundred-and-fiftieth  of  the  original  weight. 

Fig.  17  and  18  represent  in  plan  and  elevation  a  small  sampling 
mill  of  a  capacity  of  10  tons  hourly  in  use  at  a  reduction  works. 
The  process  in  this  case  is  as  follows : 

A  car,  after  weighing,  is  unloaded  into  the  storage  bins  of  the 
works,  and  a  sample  consisting  of  one-tenth  of  the  entire  carload  is 
retained  in  the  car  and  the  car  then  sent  to  the  mill.  The  sample  is 
held  upon  the  floor  until  similar  samples  from  all  the  cars  containing 
ore  of  the  same  lot  are  added  to  it.  This  sample  may  weigh  10  tons. 
It  is  now  put  through  a  Blake  crusher,  and  reduced  to  l^-in.  size. 
The  ore  from  the  crusher  then  is  elevated  and  passed  through  a 
Vezin  sampler,  while  the  rejected  portion  goes  to  one  of  the  sample- 
bins.  The  portion  to  become  the  sample,  now  weighing  4000  Ib.  or 
one-fifth  the  original  weight,  is  sent  to  a  Dodge  crusher  to  be  crushed 
to  %-in.  size.  It  is  elevated  to  a  second  Vezin  sampler,  where  one- 
fifth  is  again  cut  out,  reducing  the  amount  to  800  Ib.  It  next  passes 
to  large  rolls  that  reduce  it  to  pea-size,  then  by  an  elevator  to  the  last 
Vezin  sampler,  where  it  is  cut  to  80  Ib.  It  is  now  received  upon  the 
floor  to  be  coned  and  quartered,  and  when  reduced  to  20  Ib.  in  this 
way,  is  put  through  small  rolls  that  crush  it  to  20-mesh.  This  is 
again  reduced  in  quantity  to  5  Ib.,  put  through  a  sample-grinder, 
mixed,  quartered  to  2  Ib.,  ground  by  hand  on  a  bucking-plate  to  80- 
mesh  size,  mixed,  and  put  in  bottles  or  sample-sacks  as  in  hand- 
sampling.  The  rejected  portions  of  the  ore  are  conveyed  by  gravity 
to  any  bin  desired,  and  retained  there  until  the  ore  has  become  the 
property  of  the  reduction  works  by  purchase. 


Weight  of  ore. 

Value  in  silver  ounces  per  ton. 

Highest  300 
Average  50 

Highest  3000 
Average  75 

Highest  10,000 
Average  500 

100  tons  to  10  tons  

Cocoanut 
Orange 
Walnut 
Pea 
20-mesh 
80-mesh 

Fist 
Egg 
Chestnut 
Wheat 
25-mesh 
100-mesh 

Fist 
Walnut 
Chestnut 
Wheat 
50-mesh 
120-mesh 

10  tons  to  1  ton  

1  ton  to  200  Ib 

200  Ib.  to  5  Ib  

5  Ib.  to  bottle  sample  

Bottle  sample  . 

In  the  progressive  crushing,  as  described,  it  is  observed  that  we 


OF    THE    COMMON"    METALS.  63 

make  the  ore  finer  as  the  sample  becomes  smaller.  This  is  to  make 
sure  of  a  constant  ratio  between  the  size  of  a  single  rich  piece  and 
the  whole  sample,  that  it  may  not  produce  an  appreciable  effect  upon 
the  assay-value  of  the  sample  whether  such  a  piece  be  present  or 
absent.  The  richer,  and  more  'spotty',  or  varied  the  ore,  the  finer 
it  should  be  crushed  therefore  before  it  is  cut  or  quartered.  The 
table  above  shows  how  this  is  arranged  in  practice. 

We  may  conclude  that  for  accurate  sampling  the  principal  re- 
quirements are : 

(1)  The  taking  of  frequent  portions  to  insure  an  average  of  the 
whole  stream  that  is  undergoing  progressive  sampling. 

(2)  Thorough  mixing  of  the  ore  to  insure  uniform  richness. 
Sampling  of  ores  containing  metallic   substances. — This  is   an 

operation  requiring  a  clear  knowledge  of  the  principles  of  sampling. 
We  come  upon  these  'metallics'  sometimes  in  the  operation  of  sam- 
pling. They  must  be  separated,  cut  smaller,  and  quartered  down 
separately  by  a  hand-method,  and  reduced  in  size,  at  the  same  rate 
as  the  fine  ore.  If  fine  substance  is  made  by  cutting  up  the  metallics 
it  can  be  united  with  fine  ore.  Often  metallics  are  brittle,  but  with 
diligent  work  can  be  broken,  cut,  and  'quartered  down'  without 
serious  difficulty. 

Cost  of  sampling. — The  cost  of  moving  the  ore  cars,  unloading 
into  bins,  returning  the  cars  to  the  sampling-mill,  and  unloading 
the  fractional  part,  usually  one-tenth,  retained,  may  be  taken  at  lOc. 
per  ton.  The  cost  for  hand-sampling  the  tenth  part  may  be  taken 
at  75c.  per  ton.  Hence,  for  unloading  and  hand-sampling  a  100-ton 
lot,  the  total  cost  would  be  17.5c.  per  ton.  At  the  Metallic  Extraction 
Works,  Cyanide,  Colorado,  ore  was  unloaded  from  the  car  to  a  feed- 
shoot,  crushed  to  %-in.  size,  automatically  sampled,  and  delivered 
to  storage-bins,  for  lie.  per  ton.  A  charge  of  $1  to  $2  per  ton  has 
been  made  for  sampling,  storing,  assaying,  and  selling  ore  at  cus- 
tom works  or  sampling-mills,  where  the  company  has  acted  as  sell- 
ing agent  and  obtained  the  best  possible  price  for  the  shipper.  The 
price  for  sampling  concentrate  was  50c.  per  ton  lower. 

Sampling  concentrate  tailing,  and  ore-pulps. — Concentrate  is 
sampled  easily  for  it  can  be  thoroughly  mixed  and  sampled  by  hand. 
Tailing  contains  but  little  value,  and  needs  no  close  attention.  Ore- 
pulp  flowing  in  a  launder  is  often  automatically  sampled.  When  not 
so  sampled,  a  bucketful,  taken  each  hour  from  the  stream,  and  all 
these  samples  united  in  one  portion  after  decanting  the  water,  may 
be  used  to  determine  the  approximate  daily  average  value. 


64 


THE    METALLURGY 


13.     SAMPLING  METALS. 

Metals  may  be  sampled  either  in  the  solid  or  molten  state. 

Gold  or  silver  bars  or  ingots. — These  are  sampled  for  assay  either 
by  granulating  a  small  portion  of  them  or  by  taking  chip-samples 
from  them.  In  the  first  case,  while  the  metal  is  in  a  molten  condi- 
tion, a  small  ladelful,  weighing  an  ounce  or  less,  is  taken  from  the 
crucible  immediately  after  stirring  it.  This  is  poured  into  a  bucket- 
ful of  water  thereby  granulating  the  metal  and  forming  particles  of 
a  variety  of  sizes,  convenient  for  weighing  and  assay..  Chip-samples 
are  taken  at  points  diagonally  opposite  on  the  edges  of  the  bar,  and 
a  cold-chisel,  cutting  out  a  small  wedge-shaped  piece,  is  used  for 
this  purpose.  The  pieces  are  annealed  and  rolled  into  a  ribbon  for 
assay.  The  average  assay-value  of  the  two  pieces  thus  obtained  is 
taken  as  the  true  value. 

Base-bullion. — This  is  lead  that  comes  from  silver-lead  blast-fur- 
naces, and  it  contains  commonly  100  to  400  oz.  of  silver  per  ton. 
When  poured  into  molds  to  be  cast  in  bars,  the  silver  segregates,  and 
the  exterior  of  the  bar,  that  cools  first,  is  richer  in  silver  by  several 


2602 

Z6O.6 

264.1 

269.0 

\ 
260.9 

\ 
264.6 

250.0 

249.0 

249.0 

264.6 
1 

\~ 
263.5 

2570 

24ED 

268.0 

1 
263.6 
/ 

\ 

26/.O 

259.6 

26/.6 

/ 

Avers  ye  258.2oz.  stiver  per  Ton 

Fig.  19.     DISTRIBUTION  OF  SILVER  IN  BAR  OF  BASE-BULLION. 

ounces,  than  the  central  part.  This  is  illustrated  in  the  cross-section 
of  a  bar  (Fig.  19),  in  which  the  center  of  the  bar  assays  10  ounces 
less  than  the  exterior.  Base-bullion  is  sometimes  sampled  by  taking 
two  'chips'  or  punchings,  one  from  the  top  and  one  from  the  bottom 
of  each  bar.  The  punch  (Fig.  20) ,  resembling  a  belt  punch,  is  8  in. 
long  and  removes  a  cylindrical  piece  of  Vs-in.  diam.  by  about  1%-in. 
length. 

From  a  carload  lot  of  400  bars,  800  of  these  chips  would  be  ob- 


OF    THE    COMMON    METALS. 


tained.  These  are  melted  and  the  fused  metal  stirred  in  a  plumbago 
crucible  and  cast  into  a  bar.  This  is  a  sample  of  a  400-bar  lot  of 
base-bullion,  equivalent  to  20  tons.  A  better  way  of  sampling,  how- 
ever, is  to  re-melt  the  metal  in  a  large  kettle  (See  Fig.  172),  and  to 


Fig.    20.      BASE-BULLION   SAMPLING  PUNCH. 

skim  and  re-cast  into  bars  for  shipment.  While  casting  the  metal, 
a  sample  is  taken  from  the  molten  bath  and  poured  into  a  bullet- 
mold  of  such  a  size  that  each  bullet  weighs  approximately  a  half 
assay-ton.  This  is  trimmed  to  the  exact  weight  for  the  assay. 


,  O.34\ 


1  S00.  7\  36.  6 
O.24\  O.22\  O.22 


\/3Z./\/3S.2\/34.&/22.0\  69.  / 


O.26  O.2O 


.  /  \  672 
\0.22a230.30\0.260.20 


\  "7/.3\  7BJJ  6S.8\  6&B  • 


Fig.    21.      SECTION  OF  BAR  OF  INGOT  COPPER. 

Copper  ingots  or  anodes. — The  segregation  of  gold  and  silver  in 
copper  ingots  is  even  more  marked  than  in  bars  of  base-bullion. 
This  is  shown  in  Fig.  21,  which  represents  the  distribution  of  gold 
and  silver  in  an  ingot  of  blister-copper  5  in.  deep.  In  this  case,  how- 
ever, the  interior  is  higher  both  in  silver  and  gold.  The  usual  way  to 
sample  such  bars  is  to  drill  into  them  and  retain  the  borings  for  a 
sample.  Manifestly  a  sample  like  this  is  uncertain,  and  depends 
upon  the  selection  of  the  place  on  the  ingot  for  taking  it.  To  obviate 
this  difficulty,  in  sampling  a  lot,  say  of  100  bars,  it  is  customary  to 
drill  into  each  succeeding  bar  at  a  different  spot  to  obtain  an  average 
by  so  doing.  It  is  preferable,  however,  to  take  the  sample  at  the 
time  the  copper  is  melted  and  well  mixed  in  the  furnace  by  poling. 
As  in  the  case  of  base-bullion,  samples  of  copper  are  taken  while 


66  THE    METALLURGY 

dipping  or  casting,  one  at  the  beginning,  one  when  the  charge  is 
half  removed,  and  one  toward  the  end.  The  average  of  the  three 
samples  is  regarded  a  correct  representation. 

Pig  iron. — This  is  sampled  and  graded  by  inspecting  its  fracture 
and  by  chemical  analysis.  When  an  analysis  is  to  be  made,  the 
sample  is  taken  from  the  drillings  of  a  small  bar,  molded  while  the 
metal  is  flowing  from  the  furnace.  The  percentage  of  silicon  deter- 
mines the  grade. 

14.     CRUSHING  AND  GRINDING. 

The  method  and  extent  of  breaking,  crushing,  grinding,  or  pul- 
verizing ore  depends  upon  the  size  best  suited  to  subsequent  metal- 
lurgical treatment.  Ore  comes  from  the  mine  already  of  a  size  that 
may  need  no  further  breaking  if  intended  for  smelting.  Or  it  may 
come  to  the  reduction-works  as  concentrate,  fine  enough  for  roast- 
ing or  other  treatment.  Methods  of  breaking  or  grinding  ore  for 
further  treatment,  may  be  classified  as  follows : 

(1)  Breaking  for  the  blast-furnace. 

(2)  Breaking  for  stall  or  heap-roasting. 

(3)  Crushing  for  reverberatory  melting  or  smelting. 

(4)  Crushing  for  treatment  in  roasting-furnaces. 

(5)  Crushing  for  distillation. 

(6)  Crushing  or  grinding  for  amalgamation. 

(7)  Crushing  or  grinding  for  leaching. 

(8)  Grinding  for  sliming. 

(1)  Breaking  for  the  blast-furnace.— The  blast  ascending 
through  the  charge  carries  away  fine  ore  by  the  strong  current  and 
for  this  reason,  and  also  for  the  reason  that  coarse  ore  makes  an 
open  charge  that  permits  the  free  passage  of  the  blast,  coarse  ore  is 
preferred.  It  is  desirable,  indeed,  to  smelt  ore  as  coarse  as  it  comes 
from  the  mine,  breaking  only  the  large  lumps  with  a  hammer.  When 
the  ore  is  an  oxidized  one,  and  lumpy,  the  whole  of  it  may  be  passed 
through  a  rock-breaker  set  for  2y2-m.  crushing.  This  is  desirable 
only  when  the  blast-furnace  charge  is  too  open,  and  when  reduction, 
in  consequence,  is  poor.  Commonly  the  case  is  the  reverse  of  this 
because  of  the  large  amount  of  fine  ore  which  it  is  necessary  to  smelt. 
With  a  coarse  charge  it  is  desirable  to  crush  also  the  fluxes,  includ- 
ing iron-ore  and  limestone.  When  the  charge  is  fine,  or  tight,  the 
limestone  may  be  fed  in  50-lb.  lumps,  which  though  large  pieces, 
become  disintegrated  in  their  downward  passage  by  the  heat  that 
drives  off  the  CO2.  Up  to  this  time  the  lumps  do  duty  in  keeping 


OF    THE    COMMON    METALS.  67 

the  charge  open.  To  prevent  the  blast  carrying  away  the  finely  pul- 
verized ore,  it  is  a  good  plan  to  wet  it  before  feeding.  The  more 
effective  way  is  to  press  the  fine  material  into  bricks,  using  a  briquet- 
ting  press  (Fig.  164).  The  bricks  become  hard  on  drying  and  are 
unaffected  by  the  blast.  Briquetting  adds  about  $1  to  the  expense  of 
treatment. 

(2)  Breaking  for  stall  or  heap-roasting. — This  work  may  be 
done  either  by  hand  or  in  a  rock-breaker.    By  the  latter  method  the 
work  is  less  costly  but  more  fine  material  is  made  than  by  the  former. 
Peters,  in  his  'Modern  Copper  Smelting'  cites  a  case  of  breaking 
in  a  jaw-crusher  in  which  the  fine  resulting  amounted  to  17.3%, 
while  by  spalling  or. breaking  with  hammers,  only  9%  was  made. 
On  the  other  hand,  machine-crushing  costs  but  9c.  per  ton,  while 
hand-spalling  costs  35c.     Peters  states  that,  since  10%  fine  is  suffi- 
cient for  a  finishing-covering  for  roast-heaps  or  stalls  (See  Fig.  27 
and  28),  any  excess  above  this  quantity  should  be  avoided.    This  rule 
does  not  apply  to  fine  material  roasted  separately  in  a  reverberatory 
furnace  (See  Fig.  33). 

(3)  Crushing  for  reverberatory  melting  or  smelting. — Fig  138 
and  139  are  views  in  elevation  and  plan  of  a  furnace  in  which  ore  is 
melted  or  slagged  to  a  liquid.    Pieces  as  large  as  an  egg  melt  easily 
in  such  a  furnace.    One  great  advantage  that  the  reverberatory-fur- 
nace  has  over  the  blast-furnace  is,  that  the  damper  of  the  furnace 
may  be  closed  at  the  time  of  charging  until  the  dust,  arising  from 
the  charge  when  dropped  into  the  furnace,  subsides.    Thus  no  appre- 
ciable loss  of  fine  dusty  ore  results,  whereas  in  a  blast-furnace  such 
a  light  material  would  be  blown  away. 

(4)  Crushing  for  treatment  in  roast  ing-furnaces. — Illustrations 
and  descriptions  of  furnaces  of  this  kind  are  given  in  the  chapter  on 
oxidizing-roasting.     The  ore  must  be  crushed  fine  so  that  air  can 
reach  it  readily  to  properly  roast  it.     An  ore  consisting  mainly  of 
iron  pyrite  decrepitates  in  roasting,  so  that  it  is  fine  enough  if  crushed 
to  pass  a  2  to  3-mesh  screen.    Many  ores  and  mattes  need  crush- 
ing to  4  to  6-mesh  size.    An  ore  containing  blende  or  galena  is  com- 
pact, and  when  it  must  be  roasted,  especially  if  to  be  dead-roasted 
(that  is,  until  the  sulphur  is  eliminated),  had  better  be  ground  finer, 
or  to  10-mesh  size.     When  an  ore  is  to  be  subsequently  treated  by 
leaching,  for  which  it  should  be  more  finely  ground,  the  grinding 
can  be  done  before  roasting.    It  is  a  good  plan,  however,  to  grind  to 
a  coarse  size  for  roasting  and  re-grind  the  product  as  fine  as  de- 
sired for  the  after-treatment.     Ore  should  be  ground  no  finer  than 
needed  for  efficient  roasting,  since  more  dust  is  made  by  so  doing. 


68  THE    METALLURGY 

An  idea  of  the  efficiency  of  the  roast,  and  the  necessary  fineness 
of  the  particles  can  be  obtained  by  examining  the  latter  after  the 
roast  is  finished.  If  imperfectly  roasted,  the  particles  have  an  un- 
roasted  or  raw  core  or  center.  In  the  Stetefeldt  furnace,  the  ore  is 
ground  to  20  mesh,  since  the  time  of  roasting  is  confined  mainly  to 
the  few  seconds  occupied  by  the  fall  of  the  ore  particles  from  the 
top  of  the  tower  of  the  furnace  to  the  bottom. 

(5)  Crushing  for  distillation. — The  purpose  of  this  is  to  prepare 
ore  for  the  distillation  of  zinc  or  mercury.    When  zinc  ore  has  been 
crushed  to  8  to  14-mesh  size  it  is  fine  enough  for  the  preliminary 
roast  and  for  mixing  with  coal  in  the  subsequent  treatment.    Cinna- 
bar ore  is  fine  enough  when  1  in.  or  less  in  size.    It  is  generally  di- 
vided into  two  or  more  sizes  according  to  fineness,  each  size  being 
separately  treated  in  the  retorts. 

(6)  Crushing  or  grinding  for  amalgamation. — The  ore  is  milled 
or  crushed  in  a  stamp-battery  (See  Fig.  44  and  45).     When  com- 
minuted to  pass  the  battery-screen,  a  good  deal  of  it  is  very  fine,  or 
'slimed'   as   it  is  called.     Slime   is  not   objectionable   when   ore   is 
passed  over  the  usual  amalgamated  plate  to  collect  the  gold  it  con- 
tains.   In  silver  milling,  ore  is  first  crushed  in  a  stamp-battery,  then 
re-ground  in  pans,  thus  further  comminuting  the  coarser  particles 
that  pass  through  the  battery-screens.     Grinding  in  the  pans  is  con- 
tinued until  the  pulp  is  of  200-mesh  size  or  finer,  and  feels  smooth 
and  free  from  grit  when  rubbed  between  the  fingers. 

(7)  Crushing  or  grinding  for  leaching. — Ore  that  is  to  be  treat- 
ed by  a  leaching  method  should  be  ground  so  fine  that  the  solution 
has  access  to  the  metal  in  the  particles  of  ore.    It  must  not  be  ground 
so  fine  as  to  retard  the  leaching  by  the  slimed  material  intermingled 
with  the  sand.     In  the  operation  of  leaching  we  are  guided  by  the 
activity  or  nature  of  the  leaching  solution.    The  solution  may  contain 
chlorine,  bromine,  cyanide  of  potassium,  hyposulphite  of  soda,  com- 
mon salt,  or  consist  merely  of  water.     With  an  active  agent  like 
chlorine  the  size  may  be  less  than  with  potassium  cyanide.     Again, 
ore  may  be  porous,  or  the  precious  metal  may  be  in  a  more  soluble 
condition,  or  may  be  rendered  so  by  a  preliminary  roasting.     Thus 
Cripple  Creek  ore,  crushed  to  10-mesh  for  chlorination,  is  crushed  to 
30-mesh  size  for  cyaniding.    A  point  to  be  observed  in  preparing  ore 
for  leaching  is  to  avoid  making  slime  in  crushing.    Any  considerable 
portion   of  finely   ground   or   slimed   material  hinders   percolation 
greatly.    If  grinding  is  performed  by  rolls  a  more  granular  product, 
containing  less  slime,  is  produced  than  by  crushing  to  the   same 
screen  size  with  stamps.    This  is  shown  by  the  following  screen  an- 


OF    THE    COMMON    METALS.  69 

alysis.  The  two  products  indicated  have  passed  through  screens  of 
the  sizes  specified,  and  the  percentage  of  the  various  fine  sizes  con- 
tained are  represented  by  weight. 

TABLE  SHOWING  THE  UNDER-SIZES  RESULTING  FROM  CRUSHING. 

Wet  stamps.  Rolls. 

Crushed  through  Crushed  through 

26-mesh  screen.  23-mesh  screen. 
Screen  sizes.                                                 Per  cent.  Per  cent. 

Through  30,  on  40  mesh 11.15  9.30 

Through  40,  on  60  mesh 28.53  41.85 

Through  60,  on  90  mesh '      9.21  15.38 

Through  90 51.11  33.47 


100.00  100.00 

The  ore  used  here  is  a  conglomerate  containing  a  little  pyrite.  The 
smaller  quantity  of  slime  produced  by  the  rolls  should  be  noticed. 

If  it  is  desired  to  obtain  the  maximum  quantity  of  sand  and  the 
minimum  of  slime,  then  gradual  reduction,  or  graded  crushing, 
should  be  adopted.  This  consists  in  screening  out  the  fine  ore,  coarse- 
ly crushing  the  residue,  screening  out  its  fine  and  then  crushing 
what  is  left.  By  this  means  the  ore,  as  soon  as  crushed  to  the  de- 
sired size,  is  removed  from  further  action  of  the  crushing  machinery 
and  escapes  unnecessarily  fine  comminution. 

An  illustration  of  how  this  is  done  to  ore  that  is  to  be  ground 
sufficiently  fine  for  roasting  and  subsequent  leaching,  is  the  follow- 
ing system  of  dry-crushing.  The  assumption  here  made  is  that  the 
mill  is  to  crush  200  tons  per  day.  The  operation  is  divided  into  (A) 
coarse  and  (B)  fine  crushing. 

A.  Coarse  crushing. — A  20  by  12-in.  Blake  rock-breaker  crushes 
ore  as  it  comes  from  the  mine  containing  pieces  as  large  as  12  in. 
diam.  at  the  rate  of  25  tons  per  hour.  There  are,  however,  clayey 
and  talcose  wet  ores  containing  25  to  30%  moisture  that  stick  to  the 
rock-breaker,  and  are  impossible  to  crush  in  the  wet  state.  Such  ore 
is  first  dried  in  a  cylinderical  dryer.  The  Blake  rock-crusher  is 
shown  in  perspective  in  Fig.  22  and  in  longitudinal  section  in  Fig. 
23.  It  consists  of  a  heavy  cast-iron  frame,  marked  1,  within  which 
is  placed  the  fixed  jaw  5,  and  the  swinging  jaw  2,  and  between  them 
the  ore  is  crushed.  A  shaft  36,  eccentric  where  it  passes  through 
the  pitman  3,  causes  this  to  rise  and  fall,  producing  a  corresponding 
movement  of  the  adjacent  ends  of  the  toggles  7,  7.  As  these  rise, 
the  effect  is  to  push  the  jaw  forward  to  produce  the  crushing  move- 
ment. As  the  pitman  and  the  toggles  descend  the  jaw  recedes,  and 


70 


THE    METALLURGY 


is  pulled  back  by  the  spring  rod  16,  and  the  rubber  spring;  17.  Fly- 
wheels 11,  help  and  steady  the  movement.  The  machine  is  driven 
by  the  pulley  12,  at  250  revolutions  per  minute.  The  movement  of 
the  lower  end  of  the  jaw  is  y±  to  %  in.  For  the  breaker  above  speci- 
fied, the  discharge-opening  would  be  20  by  l1/^  in-  to  crush  to  IVo-in. 
size.  The  receiving  opening  would  be  20  by  12  in.,  and  would  take 
pieces  as  large  as  12-in.  diameter. 

The  ore,  now  crushed  to  l^-in.  size  or  finer,  goes  to  rolls  36-in. 
diam.  by  16-in.  face,  which  supply  25  tons  per  hour  to  a  revolving 


Fig.    22.      THE  BLAKE  ORE-CRUSHER. 


screen  or  trommel  having  %-in.  holes.  A  set,  or  pair,  of  such  rolls 
of  the  belt-driven  type  is  shown  in  Fig.  24.  One  roll  is  carried  by  its 
shaft  in  fixed  bearings  or  boxes,  and  is  driven  by  the  large  pulley. 
The  other  roll,  called  the  movable  one,  is  held  to  its  work  by  power- 
ful spiral  springs  acting  on  the  shaft-boxes  with  a  pressure  of  15 
tons  upon  each  box.  Thus  the  pressure  of  the  springs  is  such  that 
if  a  hard  object,  such  as  a  hammer-head,  were  to  fall  between  the 
rolls,  they  would  open  only  under  a  pressure  of  30  tons.  The  smaller 
pulley  on  the  shaft  of  the  movable  roll  is  intended  to  keep  the  roll 
in  motion  at  the  speed  of  the  other,  but  not  for  other  work,  the 


OF    THE    COMMON    METALS. 


71 


power  being  mostly  transmitted  through  the  large  pulley.  The  rolls 
(in  Fig.  24),  are  covered  by  the  housing.  The  feed-hopper  of  the 
rolls  is  seen  immediately  above  the  housing.  As  the  ore  drops  from 
the  rolls  it  falls  through  a  chute  to  the  boot  of  a  belt  or  chain-eleva- 
tor (See  Fig.  189  and  191). 

The  material  from  the  rolls  is  delivered  or  discharged  through 
a  chute  into  a  Vezin  sampling-machine  (Fig.  16),  making  40  revolu- 
tions per  minute.  This  takes  out  one-fifth  for  a  sample,  and  leaves 
four-fifths  to  go  to  the  storage-bins.  The  sampling  is  completed  by 


Fig.    23.      BLAKE  ORE-CRUSHER   (SECTION). 

the  methods  described  under  'machine-sampling'.  A  single  samp- 
ling-machine can  easily  handle  the  200  tons  crushed  during  the  10- 
hour  day  shift. 

The  ore  %  to  %  in-  and  less  in  size,  is  drawn  as  desired  from  the 
storage-bins  into  two-wheeled  buggies,  and  dumped  into  the  hopper 
of  the  cylindrical  drier.  A  feeding-shoe  or  some  other  kind  of  au- 
tomatic feeder,  supplies  it  continuously  to  the  drier,  shown  also  at  /, 
Fig.  110  and  111.  This  drier  is  24  ft.  long  and  of  sufficient  capacity 
to  dry  ore  containing  6%  moisture  to  \%  or  less,  at  the  rate  of  10  to 
15  tons  per  hour,  heating  it  at  the  same  time  so  that  it  will  screen 


72  THE    METALLURGY 

readily.  When  in  this  condition  the  ore  is  'lively',  and  will  screen 
without  difficulty,  but  if  it  were  not  for  the  drying,  it  would  stick 
to  the  screen  and  clog  it.  The  cost  of  this  drying  approximates  5c. 
per  ton. 

B.  Fine  crushing. — This  work  is  done  continuously.  Fig.  25  is 
a  diagram  showing  the  series-crushing  system  used.  The  ore-supply 
from  the  drier  goes  to  the  roughing-rolls  a  which  reduce  it  from  0.75 
to  0.25  inch.  The  crushed  ore  is  raised  by  the  elevator  to  a  trom- 
mel, or  separating-screen,  having  screens  of  Vs  and  %-in.  aperture, 
respectively.  The  first  two-thirds  of  the  screen  takes  out  all  ma- 
terial less  than  %  in.,  and  the  final  size  is  all  coarser  than  y2  inch. 


Fig.   24.      CRUSHING  ROLLS. 

We  thus  get  three  products,  an  oversize  from  the  coarser  screen 
which  goes  back  to  the  roughing  rolls  to  be  re-crushed ;  a  screened 
product  or  under-size,  which  goes  to  the  medium  rolls  &,  set  at  %  in-> 
there  to  be  crushed  and  sent  back  to  the  separating-screen ;  and  fin- 
ally, an  under-size  through  the  %-in.  screen,  fine  enough  to  go  to 
the  finishing  rolls.  Until  crushed  so  fine  that  it  passes  the  finest 
mesh,  the  ore  is  returned  to  the  trommel. 

The  fine  product  of  the  screen  is  raised  by  the  elevator  /'  and  the 
ore-stream  is  equally  divided  between  the  finishing  screens  s"  and 
s"'  which  are  provided  with  30-mesh  wire  cloth.  The  undersize 
from  these  revolving  screens  or  trommels  drops  into  the  storage  bin 
m,  while  the  oversize  is  conveyed  to  the  finishing  rolls,  after  which 
it  goes  again  by  elevator  to  the  finishing  screens.  Thus  nothing  en- 


OF    THE    COMMON    METALS. 


73 


ters  the  storage-bin  except  30-mesh  or  finely  ground  ore    (0.02-in. 
diam.),  ready  for  further  treatment. 

We  observe  that  any  8-mesh  or  finer  product,  that  has  been  pro- 
duced by  roughing  rolls,  does  not  go  to  the  medium  rolls,  and  that 
any  finished  sand,  in  the  8-mesh  product,  is  not  forced  upon  the 


Fig.    25-.      FLOW-SHEET  FOR  DRY  CRUSHING. 

finishing  rolls,  but  goes  first  to  the  finishing  screens.  Thus  the  mo- 
ment a  particle  is  broken  to  a  finished  condition,  it  passes  to  the 
finished-product  bin  without  passing  through  the  crushing  machin- 
ery. One  square  foot  of  screen  will  separate  6  cu.  ft.  of  product  per 
24  hours.  The  200  tons  required  can  be  treated  in  20  to  22  hours, 
allowing  time  for  delays  and  repairs.  This  gives  a  capacity  for  each 


74  THE    METALLURGY 

set  of  finishing  rolls  of  5  tons  per  hour.  The  speed  of  the  rolls  varies 
with  the  coarseness  of  the  material  to  be  crushed.  The  coarse  rolls, 
run  at  600-ft.  the  fine  ones  at  1000  to  1400-ft.  peripheral  speed,  per 
minute. 

This  system  of  graded  crushing  is  preferable  because  the  pro- 
duct contains  a  minimum  of  fine  or  slimed  ore ;  and  being  granular, 
is  more  easily  percolated  or  leached.  As  each  piece  or  particle  of 
ore  is  crushed  by  a  single  nip,  the  fine  is  separated  by  the  screen  and 
protected  from  unnecessary  breaking  with  consequent  waste  of 
power.  The  chief  costs  in  this  system  of  dry-crushing,  are  those  of 
labor,  power,  supplies,  and  repairs.  These  vary  with  the  tonnage. 


Fig.    26.      TROMMEL  OR  CYLINDRICAL  REVOLVING  SCREEN. 

The  cost  of  crushing  to  30-mesh  size,  in  preparation  for  roastirg  ;>L* 

leaching,  is  as  follows : 

For  coarse-crushing  and  automatic  sampling. .  .$0.106 

For  drying  and  fine-crushing    0.275 

For  power    0.105 


$0.486 

In  round  numbers  the  cost  is  50c.  per  ton,  but  to  this  must  be  added 
general  expense,  including  management,  office  expense,  rates,  taxes, 
insurance,  cost  of  water,  and  improvements. 

Besides  crushing  with  rolls,  after  coarse-crushing,  ore  may  be 
crushed  by  stamps  or  Chilean  mills.  Fig.  44  is  a  stamp-battery,  a 
favorite  and  efficient  machine;  Fig.  100  is  a  Chilean  mill.  This  will 
be  described  later. 

(8)  Grinding  for  sliming. — Ore  frequently  contains  gold  and 
especially  silver  so  finely  desseminated  in  the  substance  of  the  ore 
that  giinding  finer  than  150-mesh  is  needed  to  unlock  the  minerals 
from  the  adhering  gangue.  While  rolls  or  stamps  grind  ore  to  a 


OF    THK    COMMON    METALS.  75 

certain  fineness,  further  grinding,  after  water-separation  or  classi- 
fication of  sand  and  slime,  is  performed  upon  the  granular  or  sandy 
portion  too  coarse  to  pass  uncrushed.  Grinding-pans  and  tube-mills 
have  been  used  for  this  purpose,  and  have  proved  most  satisfactory. 
The  tube-mill  (Fig.  96),  later  to  be  described,  is  a  favorite  means 
employed  for  this  purpose  because  of  its  great  capacity,  simplicity, 
and  cheapness  in  repairs.  The  cylinder  or  tube  revolves  20  times 
per  minute,  and  half-filled  with  pebbles  the  size  of  the  fist  rolling 
over  one  another,  effectually  comminutes  the  ore  particles.  Grind- 
ing to  any  desired  size,  is  done,  and  the  coarser  particles  which 
escape  grinding,  are  classified  and  returned  to  the  machine  for  re- 
grinding. 


PART  II.     ROASTING 


PART  II.     ROASTING. 

15.     OXIDIZING-ROASTING. 

This  operation  is  also  called  *  calcining',  though  the  term  calcin- 
ing is  applied  also  to  heating  carbonates  to  expel  carbon  dioxide. 
We  shall  proceed  to  the  chemistry  of  the  process. 

16.     THE  CHEMISTRY  OF  ROASTING. 

The  operation  of  roasting  is  for  the  purpose  of  burning  or  expell- 
ing the  sulphur,  which  the  ore  contains,  by  the  action  of  heat  with 
the  access  of  air.  This  can  be  done  either  when  ore  is  in  lump-form, 
in  heaps,  or  when  fine,  by  means  of  reverberatory  furnaces.  The 
various  types  of  these  furnaces  are  to  be  described  and  illustrated  in 
the  present  chapter. 

For  the  purpose  of  illustration,  let  us  consider  the  action  that 
takes  place  in  a  reverberatory  furnace  such  as  is  shown  in  Fig.  31, 
32,  and  33.  In  this  furnace  the  fire  is  at  the  end,  and  the  charge  is 
spread  upon  the  hearth.  As  roasting  proceeds,  the  roasted  ore  at  the 
fire-end  is  withdrawn  and  the  part  remaining  unroasted  is  moved 
from  time  to  time  toward  the  fire,  fresh  ore  being  constantly  sup- 
plied at  the  upper  end.  After  20  hours,  ore  fed  at  the  cool  end,  is  in 
the  hottest  part  of  the  furnace.  Let  us  take  for  example  an  ore  with 
a  silicious  gangue,  containing  mixed  sulphides  such  as  galena,  blende, 
pyrite,  and  chalcopyrite.  The  silica  tends  to  prevent  sintering  and 
makes  the  charge  more  open  or  accessible  to  the  air.  To  do  good 
roasting,  certain  requirements  must  be  observed.  First,  the  heat 
must  be  sufficient  to  start  the  roasting.  Later,  the  temperature  must 
be  high  to  drive  off  the  last  of  the  sulphur.  Air  must  be  abundant, 
to  oxidize  the  ore  freely,  and  to  carry  away  readily  the  products  of 
the  re-action.  The  surface  exposed  to  the  heat  and  air  should  be  ex- 
tensive. Finally,  the  ore  should  be  stirred  frequently,  to  bring  fresh 
surfaces  into  contact  with  the  air. 

The  ore  is  dropped  upon  the  hearth  at  the  cool  end  of  the  furnace, 
where  the  temperature  (350°C.)  is  sufficient  to  expel  moisture  and 
start  the  reaction  of  combustion.  In  10  to  15  minutes  the  charge 


80  THE    METALLURGY 

becomes  hot  enough  for  the  oxidation  of  pyrite  to  begin,  as  evinced 
by  the  blue  flickering  flame  that  plays  over  the  surface  of  the  charge. 
Pyrite,  under  the  action  of  heat,  separates  thus : 

(1)  FeS2  +  heat  +  20  =  FeS  +  S02. 

The  sulphur  expelled  combines  with  oxygen  from  the  air.  An  equiva- 
lent, or  32  Ib.  of  sulphur,  burning  to  S02,  yields  72,000  calories,  or 
72,000  -r-  32  =  3220  pound-calories  per  pound  of  sulphur  burned. 
The  S02  is  removed  by  the  draft.  The  remaining  FeS  and  also  CuS, 
the  ZnS,  and  PbS,  begin  to  oxidize,  the  activity  of  roasting  being  in 
the  order  here  named.  While  oxidation  proceeds  with  them  all,  the 
FeS  is  most  easily  roasted,  while  the  ZnS  and  PbS  are  the  slowest  in 
parting  with  sulphur.  Beginning,  therefore,  with  the  FeS  we  have : 

(2)  FeS  +  30  =  FeO  +  SO2 

23,800  66,400     71,000  =  113,600; 

or  in  words  the  iron  sulphide  becomes  oxidized  to  ferrous  oxide,  with 
the  evolution  of  S02,  the  reaction  being  accompanied  by  heat  to  the 
extent  of  113,600  -r-  32  =  3550  pound-calories  per  pound  sulphur 
burned,  as  explained  in  Part  I,  Section  3,  under  '  Thermo-chemistry '. 
The  cupric  sulphide  of  the  chalcopyrite  acts  according  to  the  for- 
mula : 

(3)  CuS  +  30  =  CuO  +  S02 

10,200  37,200     71,000  =  +  98,000 ; 

or,  per  pound  sulphur,  98,000  -f-  32  =  3030  calories.  The  blende,  un- 
der the  action  of  the  heat  and  air,  is  affected  in  the  same  way : 

(4)  zns  -f  30  =  ZnO  +  S02 

43,000  86,400     71,000  =  +  104,200 ; 

or,  per  pound  sulphur  present,  3420  calories.  Galena  also  roasts  ac- 
cording to  the  reaction : 

(5)  PbS  +  30  =  PbO  +  S02 

17,800  51,000     71,000  =  +  104,200 ; 

which  gives  us  3250  calories  per  pound  sulphur.  It  will  be  noticed 
that  the  heat  evolved  per  pound  sulphur  burned  is  much  the  same  in 
each  case ;  and  hence  the  sulphide,  containing  most  sulphur,  gives  off 
the  most  heat  in  roasting.  These  re-actions,  especially  of  blende  and 
galena,  proceed  gradually  throughout  the  roasting  operation.  The 
air  acts  chiefly  on  the  exposed  surface,  and  the  activity  of  the  above 
reactions  is  accordingly  increased  by  stirring  the  charge.  The  FeS 
that  has  been  formed,  as  shown  in  reaction  (1),  and  which  lies  upon 
the  surface,  is  exposed  to  an  excess  of  air.  The  FeS,  in  presence  of 
the  silica  of  the  ore,  acts  by  catalysis  and  is  a  sort  of  go-between  for 


OF    THE    COMMON    METALS.  81 

the  FeS2  and  the  oxygen  of  the  air.    It  finally  becomes  oxidized  thus : 

(6)  3FeS  -f  HO  =  2SO2  +  Fe,03  +  FeS04 

3x23,800  2x71,000    199,400    235,600  =  +  505.600 ; 

or,  per  pound  sulphur,  5260  calories.  This  is  observed  to  be  a  most 
energetic  exothermic  reaction.  The  products  of  the  reaction  are 
iron  sulphate,  SO2  which  is  carried  away  by  the  draft,  and  Fe2O3 
which  is  acted  upon  by  more  FeS  when  the  two  are  stirred  together 
as  follows : 

(7)  FeS  +  10Fe203  =  7Fe304  +  S02 

23,800  10x199,400  7x265,800  71,000  =  86,200; 

If  the  charge  is  not  stirred,  reaction  (6)  causes  a  red-colored  pro- 
duct to  form  on  the  surface  of  the  charge.  This  is  the  compound, 
Fe2O3,  but  on  stirring  it  decomposes  into  Fe3O4  as  shown  in  equation 
(7),  and  the  product  becomes  black.  When  roasted  to  excess,  the 
product  takes  on  this  red  color,  undesirable  in  some  roasting  opera- 
tions. 

As  the  ore  is  moved  to  a  hotter  part  of  the  furnace,  the  activity 
of  the  above  reactions  continue,  and  at  an  incipient  red.  (590°C.), 
iron  sulphate,  which  has  been  formed  in  the^mfmnfSr  tfoffAn/egins  to 
decompose,  reacting  on  the  cupric  oxide  as  follows. 


Thus,  as  the  SO3  is  given  off  frojnr^the  FeS04,  and  while  in  afn  as- 
cent condition,  it  is  taken  by  the  CuO,  the  reaction  being  an  endo- 
thermic  one.  Sulphate  'Li  funiiGd  Hum* 

(9)     FeS04  +  2CuO  =  FeO  +  CuSO4 

235,600      37,200    66,400     182,600=     -23,800. 

This  reaction  is  t-n  un  mui'ivendothermic  iLan  lliu  fui'moj,,  and  hence 
is  only  a  partial  one.  There  remains  uncombined  some  iron  sulphate, 
which  is  decomposed  by  the  heat  thus : 

(10)     FeSO4  =  FeO  +  SO3 

235,600     66,400    91,800=  -77,400. 
This  also  is  an  endothermic  reaction. 

At  a  slightly  greater  heat  (655°C.),  the  cupric  sulphate,  formed 
but  a  short  time  previous  according  to  reaction  (8),  begins  to  de- 
compose, and  at  a  dull  red  heat  (705°C.),  the  corresponding  decom- 
position of  the  cupro-cupric  sulphate  begins.  These  decomposition- 
reactions  of  the  copper  sulphate  are  complete  at  850°C.,  or  at  a 
cherry-red  heat.  To  this  point,  they  indicate  the  action  that  is  going 


82  THE    METALLURGY 

on  in  a  sulphatizing  roast  like  that  of  the  Ziervogel  process  to  be 
later  described. 

At  850°C.  the  zinc  and  lead  oxides,  reacting  on  the  copper  sul- 
phate that  is  being  decomposed  by  the  heat,  begin  to  be  changed,  to 
sulphates. 

(11)  ZnO  +  CuS04  =  ZnS04  +  CuO 

86,400    182,600     230,000     37,200=  -1800. 

(12)  PbO  +  CuS04  =  PbS04  +  CuO 

51,000     132,600     216,200     37,200=  -19,800. 

In  general  we  may  note  that  reactions  (8)  to  (12),  inclusive,  are  eri- 
dothermic,  and  in  place  of  helping  along  the  roasting  by  heat  de- 
veloped, as  shown  in  the  early  action  of  the  charge,  demand  heat  and 
absorb  it  as  the  result  of  the  reactions.  Fortunately  these  later  re- 
actions take  place  only  at  the  hot  end  of  the  furnace. 

As  the  charge  is  moved  nearer  the  fire  the  sulphates  completely 
decompose,  the  zinc  sulphate  doing  so  more  readily  than  the  lead 
sulphate.  At  1050°0.  (a  dark-orange  heat),  copper  oxide  is  decom- 
posed to  cuprous  oxide,  and  ferric  oxide  to  Fe304,  oxygen  escaping 
at  this  high  temperature. 

At  this  stage  the  ore  begins  to  fuse,  if  it  contains  lead,  but  when 
little  lead  is  present,  it  only  slightly  agglomerates.  If  the  ore  con- 
tained much  leadj  even  this  temperature  can  not  be  attained  without 
causing  it  to  begin  to  slag.  In  that  case  the  charge,  rendered  no 
longer  porous  or  accessible  to  the  air,  necessarily  ceases  to  roast,  zinc 
and  lead  sulphates  decompose  imperfectly,  and  the  sulphur  is  not 
well  eliminated.  It  is  hard  to  roast  a  leady  ore  well.  On  the  other 
hand  a  zinc  ore,  free  from  lead,  can  and  should  be  brought  to  a  high 
finishing-heat  when  it  is  desired  to  decompose  the  zinc  sulphate  and 
remove  the  last  portions  of  the  sulphur. 

Sometimes,  with  lead-bearing  zinciferous  ores,  to  be  treated  in  the 
silver-lead  blast-furnace,  after  roasting,  the  roaster  is  arranged  with 
a  fuse-box  as  shown  in  Fig.  33.  The  charge,  which  has  been  roasted 
on  the  roasting-hearth  of  the  main  part  of  the  furnace,  is  moved 
from  the  hearth  into  the  fuse-box,  and  melted  after  adding  some  sil- 
icious  ore.  The  silica  reacts  on  the  zinc  and  lead  sulphates  thus: 

(13)  ZnS04  +  Si02  =  ZnSiO3  +  SO3  and 

(14)  PbSO4  +  Si02  =  PbSi03  +  S03 

The  sulphur  is  eliminated  as  sulphuric  anhydride  fume,  and  the  re- 
sulting product  is  freed  from  sulphur.  Zinc  silicate  enters  slag  as 
such,  and  is  eliminated  by  that  means.  Lead  silicate  is  reduced  in  the 
blast-furnace  with  the  recovery  of  the  lead.  The  charge,  originally 


OF    THE    COMMON    METALS.  83 

pulverous,  is  now  in  the  form  of  a  slag,  so  that  no  loss  of  flue-dust 
results. 

It  has  been  found  that  with  2%  S02  by  volume,  or  4.4%  by 
weight,  in  the  escaping  gas,  roasting  is  active.  This  corresponds  to 
23  Ib.  air  per  pound  S02  or  46  Ib.  (570  cu.  ft.)  air  per  pound  sulphur 
driven  off.  Calculating  this  for  a  16-ft.  McDougall  roasting-furnace, 
treating  40  tons  of  ore  per  24  hours,  and  roasting  it  from  35%  down 
to  7%  sulphur  in  the  product,  we  have  an  elimination  of  approxi- 
mately i/4  Ib.  of  sulphur  per  second,  which  needs  142  cu.  ft.  free  air, 
equal  to  284  cu.  ft.  of  escaping  gas  at  273 °C.,  a  temperature  which 
gives  a  maximum  chimney-discharge  as  explained  in  the  chapter  on 
combustion.  For  a  velocity  in  the  stack  and  flues  of  20  ft.  per  sec- 
ond as  a  maximum,  this  would  require  an  area  of  14.2  sq.  ft.,  or  a  di- 
ameter, for  a  round  stack,  of  4  feet  3  inches. 

When  more  than  46  cu.  ft.  of  air  per  pound  of  sulphur  is  ad- 
mitted, though  roasting  proceeds  actively,  owing  to  a  good  supply  of 
fresh  air  and  to  the  fact  that  the  products  of  ihe  reactions  are  re- 
moved promptly,  still  the  furnace  becomes  cooled  too  much.  If  less 
air  than  this  is  admitted,  roasting  proceeds  more  slowly  so  that  with 
4%  S02  by  volume  the  roasting  is  slow,  with  8%  very  slow,  and 
with  9%  it  ceases  entirely. 

The  various  reactions  described  above  need  time ;  and  the  larger 
the  body  of  ore,  the  longer  the  time  must  be  to  complete  the  roast. 
If  a  few  grams  of  ore  are  roasted  in  the  muffle,  the  operation  is  com- 
plete in  half  an  hour,  but  in  a  hand-reverberatory  furnace  (like  that 
shown  in  Fig.  32),  roasting  14  tons  per  day,  the  operation  takes  20 
hours. 

The  temperatures  at  which  the  reactions  described  above  take 
place  are  as  follows : 

At  150 °C.  the  odor,  due  to  the  volatilization  of  some  of  the  first  or 
loosely-held  sulphur  in  pyrite,  can  be  detected. 

At  350°C.  the  sulphur  of  the  sulphides,  particularly  of  the  iron 
sulphide,  begin  to  burn. 

At  590 °C.  the  iron  sulphate,  formed  at  a  lower  temperature,  be- 
gins to  decompose. 

At  655 °C.  copper  sulphate  (CuSOJ  (see  equation  8),  decomposes. 

At  705°C.  cupro-cupric  sulphate  (CuO,SO4),  formed  at  the  time 
copper  sulphate  is  formed  (see  equation  9),  begins  to  decompose. 

At  850°C.  copper  sulphates  are  entirely  decomposed. 

At  835  to  850 °C.  the  maximum  amount  of  soluble  silver  sulphate, 
(AgSOJ  (when  silver  is  present  in  the  ore)  is  formed. 

At  1050°C.  copper  oxide  (CuO)  is  decomposed  to  Cu2O. 


84  THE    METALLURGY 

At  1100°C.  ferric  oxide  (Fe,03)  is  decomposed  to  the  next  lower 
oxide  Fe3O4. 

Reactions  in  blast-roasting  copper-bearing  sulphides. — When  a 
mixture  of  iron  and  copper  sulphide  is  pot-roasted,  desulphurization 
proceeds  rapidly  if  the  ore  be  wet  and  silica  added ;  otherwise  it  pro- 
ceeds slowly.  For  these  reactions  we  have : 

(15)  3FeS  +  4H2O  =  Fe304  +  3H2S  +  2H, 

(16)  2Fe,O3  -f  7H2S  =  4FeS  +  3S02  +  14H. 

When  air  is  blown  into  the  charge  both  hydrogen  and  H2S  burn.  In 
a  well-burned  charge  some  Fe203  can  be  seen  at  the  sides  and  top 
where  cooled  by  radiation,  but  Fe304  reacting  on  FeS  gives  FeO  as 
follows : 

(17)  FeS  +  3Fe304  ==  lOFeO  +  SO2. 

This  reaction  is  exothermic  and  at  a  high  temperature  with  silica 
would  form  ferrous  silicate,  again  producing  heat.  Indeed,  in  action 
the.  formation  of  this,  with  the  consequent  sintering,  can  be  seen 
spreading  as  the  burning  proceeds. 

17.     ROASTING  ORBS  IN  LUMP  FORM. 

Heap-roasting. — Roasting  in  heaps  sulphide  ores  containing  cop- 
per is  an  operation  that  must  be  performed  with  knowledge  and 
care,  in  order  to  obtain  satisfactory  results.  Before  selecting  this 
method  a  careful  study  of  the  environment  must  be  made.  It  can 
not  be  adopted  in  a  settled  country  where  the  fumes  would  be  a  nuis- 
ance, or  where  it  would  be  likely  to  injure  live-stock,  or  vegetation, 
or  damage  crops.  In  the  arid  and  scarcely  settled  Rocky  Mountain 
region,  of  the  United  States  and  Mexico,  it  may  be  employed  to  ad- 
vantage. Around  an  installation  of  moderate  size,  where  no  more 
than  25  tons  of  sulphur  escape  into  the  atmosphere  daily,  the  area 
affected  may  be  not  more  than  four  miles  in  extent ;  and  the  site  for 
the  roasting-plant  may  be  chosen  with  this  in  view.  Often  the  pre- 
vailing winds  blow  from  a  single  quarter,  and  permit  placing  the 
roast-piles  where  they  give  little  offence.  The  location  should  be 
chosen  with  regard  to  the  needs  of  the  plant  itself,  so  that  smoke 
and  fume  will  seldom  be  driven  into  the  buildings. 

A  roast-yard  should  be  of  ample  size,  approximately  level,  and  so 
drained  that  the  surface  water  shall  not  flow  over  it.  At  Jerome, 
Arizona,  a  track  follows  the  contour  of  the  hillside,  and  the  roast- 
heaps  are  ranged  close  beside  it  so  that  the  ore  is  trammed  conveni- 
ently to  the  pile,  unloade<J  upon  it,  and  roasted.  After  roasting  the 
ore  is  loaded  at  these  piles,  without  waste  of  labor,  and  is  trammed 


OF    THE    COMMON    METALS. 


85 


to  the  smelter.  The  roast-heaps  are  so  far  apart  that  the  fume  from 
one  pile  does  not  interfere  with  making  another. 

A  roast-pile  40  ft.  wide  and  6  ft.  high,  containing  240  tons,  will 
burn  70  days.  To  this  time  must  be  added  10  days  for  removing 
and  re-building.  This  equals  three  tons  of  ore  roasted  per  day. 

In  preparing  a  roast-pile,  a  foundation  is  first  provided,  by  level- 
ing the  ground  and  making  a  surface  of  clayey  loam.  Upon  this  is 
placed  a  layer  of  fine  ore  3  or  4  in.  thick  upon  which  the  pile  is  to 
be  made.  This  layer  gradually  roasts,  and  as  fast  as  this  occurs,  it 
is  removed  and  sent  to  the  furnaces,  and  fresh  fine  is  used  in  its 
place.  The  ore  is  often  hauled  to  the  ore-heap  in  carts,  or  brought 
in  wheelbarrows,  or  conveyed  along  a  system  of  tracks  from  the 
higher  ground  adjoining,  or  from  the  mine.  This  is  shown  in  Fig. 


Fig.  27.      ROAST-YARD  WITH  TRESTLE    (CROSS  SECTION). 


Fig.    28.      ROAST-YARD  WITH  TRESTLE    (LONGITUDINAL  SECTION). 

27  and  28.  The  bents,  as  shown  in  Fig.  28,  are  36  ft.  apart,  having 
trussed  stringers  10  by  12  in.  to  support  the  tracks  over  the  roast- 
•heaps.  In  Fig.  28  a  completed  head  is  shown  at  the  right,  in  the 
middle  is  the  cross-section  of  another  heap,  and  at  the  left  one  just 
started.  Fig.  27  represents  a  section  of  a  heap  and  the  trac.k  that 
runs  past  every  heap,  by  which  the  roasted  ore  is  taken  away.  In 
constructing  the  heap,  a  turn-plate,  similar  to  that  shown  in  Fig.  30, 
is  placed  across  the  track  at  the  desired  point,  and  a  temporary  track 
of  stout  rails,  supported  on  temporary  trestles,  run  out  from  the 
main  track  at  right-angles. 

The  required  height  of  the  pile  depends  upon  the  character  of 
the  ore.  An  ore  of  15%  sulphur  may  be  made  into  a  pile  9  ft.  high, 
while  massive  pyrite  requires  but  6  ft.  for  the  best  conditions  and 


86  THE    METALLURGY 

thorough  roasting.  Upon  the  bottom,  above  described,  is  placed  the 
fuel  for  starting  the  burning.  This  consists  of  a  layer  of  wood  4  to 
8  in.  thick,  a  4-in.  layer  being  sufficient  for  the  high-sulphide  ores. 
The  wood  may  be  of  any  kind  and  length,  where  ore  rich  in  sulphide 
is  to  be  roasted.  Cheaper  wood,  such  as  crooked  branches,  logs,  and 
old  tree-trunks  can  be  used  toward  the  center.  The  best  pieces 
should  be  selected  for  the  borders.  The  interstices  between  these 
closely-packed  pieces  may  be  closed  by  placing  in  them  fine  sticks, 
brush-wood,  or  chips,  so  that  the  ore  will  not  fall  through.  Three 
chimneys  8  in.  square,  made  of  four  old  boards,  nailed  together,  are 
set  upright  on  the  layer,  and  connected  to  channels  below  to  create 
a  draft  through  the  wood  when  the  pile  has  been  fired.  In  Fig.  28 
one  of  these  three  chimneys  is  shown.  They  also  can  be  made  of 
round  sticks  wired  together,  or  of  waste  sheet-iron,  bent  in  cylind- 
rical form. 

It  must  be  noted  here  that  often  too  much  wood  is  used.  About 
one  cord  is  required  for  40  tons  of  ore.  It  must  be  remembered  that 
the  wood  is  merely  to  start  the  burning  of  the  pile.  It  soon  burns 
out,  but  the  ore  continues  to  burn  by  its  own  action.  The  more  mas- 
sive or  sulphurized  the  pyrite,  the  less  wood  is  needed.  The  pile 
should  be  kept  burning  uniformly,  and  no  attempt  should  be  made 
to  hasten  the  process  unduly,  since,  with  an  excessive  heat,  the  ore 
fuses,  causing  the  action  to  stop.  The  coarse  ore,  broken  so  that  no 
pieces  are  more  than  4  in.  diam.,  and  varying  from  that  size  down 
to  1  in.  diam.,  is  dumped  upon  the  layer  of  wood.  The  pile,  built  to 
the  required  height  with  as  steep  an  angle  at  the  sides  as  will  stand, 
forms  a  shapely  frustrum  of  a  pyramid  with  sharp  corners.  Upon 
this,  as  shown  in  both  figures,  is  placed  ragging  (made  by  screening 
out  all  the  ore  between  one  inch  and  one-fourth  inch  in  size),  form- 
ing a  layer  over  the  top  and  at  the  sides,  thicker  at  the  bottom  and 
thinner  toward  the  top  of  the  slopes.  Outside  this  is  spread  a  layer 
of  fine  ore  of  less  than  one-fourth  inch  diameter  screened  from  the 
ragging,  making  a  thin  coating  over  the  exterior  surface. 

The  pile  is  fired  at  the  different  channels  around  the  edges,  put- 
ting in  some  kindling  wood  to  start  combustion  and  selecting  a  time 
of  quiet  weather.  After  4  to  6  hours,  the  fire  having  become  well 
distributed  through  the  pile,  more  fine  ore  is  put  on  the  steep  slop- 
ing sides  of  the  pile,  thin  above  and  thick  below.  The  heap,  especi- 
ally at  first,  must  be  closely  watched,  fine  being  applied  to  check  the 
draft  where  too  vigorous,  or  holes  being  opened  in  the  covering  to 
draw  the  fire  in  the  direction  of  places  where  it  is  not  active.  Fine 
is  used  freely  to  control  the  fire,  throwing  it  on  with  a  shovel 


OF    THE    COMMON    METALS.  87 

wherever  needed,  along  the  borders,  at  the  sides  and  top  of  the 
heap.  One  may  go  upon  the  heap  to  regulate  the  combustion,  by 
taking  advantage  of  the  wind  blowing  the  smoke  to  one  side. 

Moderate  rains  and  snow  have  but  little  effect  upon  the  roasting ; 
a  high  wind  is  apt  to  stop  the  burning  on  the  windward  side.  This 
may  be  prevented  by  a  temporary  fence  placed  as  a  wind-break. 
Abundant  rains  tend  to  leach  out  and  wash  away  copper  sulphate. 
Where  such  conditions  exist  the  heaps  may  be  roofed  over,  or  where 
this  is  thought  to  be  too  great  an  expense,  some  saving  can  be  made 
by  draining  through  ditches  to  a  launder  containing  scrap-iron,  upon 
which  the  copper  precipitates.  In  the  dry  region  of  the  Western 
United  States,  rains  are  moderate,  and  there  are  long,  dry  periods. 
In  countries  like  Mexico,  that  have  a  rainy  season,  heap-roasting  can 
be  suspended  until  the  rains  are  over. 

Wherever  possible  roast-heaps  should  be  left  undisturbed  until 
the  completion  of  the  roast,  but  if  ore  is  badly  needed,  it  may  be 
taken  from  the  roasted  and  cooled  portions  of  the  pile,  without  dis- 
turbing the  action  of  the  hot  core,  still  burning.  The  best  way  is  to 
start  roasting  operations  several  weeks  in  advance  of  the  probable 
need  of  the  smelter,\  so  as  to  have  available  a  sufficient  supply  of 
well-roasted  ore.  With  care  it  should  be  possible  to  roast  90%  of 
the  ore  in  the  pile,  including  the  fine.  When  removing  the  ore,  any 
unroasted  pieces,  which  feel  heavier  than  the  well-roasted  ore,  can 
be  sorted  out  and  thrown  upon  the  next  roast-heap. 

Heap-roasting  of  matte. — Matte  can  be  well  roasted  in  lump 
form,  but  unlike  ore,  it  requires  two  or  more  burnings.  After  the 
first  firing,  in  spite  of  care,  matte  shows  but  little  the  change  it  has 
undergone.  At  the  second  burning,  using  a  larger  quantity  of  wood, 
the  result  of  the  first  burning  begins  to  show.  A  large  portion  of 
the  twice-burned  material  is  found  to  be  light  in  weight  and  porous, 
and  to  contain  no  unburned  core.  In  fact  the  thoroughness  of  the 
roast  may  be  judged  by  feeling  of  the  lumps  with  the  hand.  If  well 
roasted  lumps  are  broken,  they  no  longer  show  the  raw  core  at  the 
center. 

The  bed  of  wood  can  be  prepared  for  matte  as  for  ore,  but  the 
pile  is  smaller,  being  only  12  ft.  square  by  6  ft.  deep,  with  a  single 
chimney  at  the  center.  The  broken  matte,  with  the  raw  fine  spread 
over  it,  is  covered  with  the  finer  portion  of  the  roasted  ore.  The 
burning  of  the  heap  lasts  11  days,  and  when  ended,  the  heap  is  taken 
down,  and  the  imperfectly  roasted  part  made  into  a  new  pile,  and 
the  roasted  matte  sent  to  the  furnace.  It  is  a  good  plan  in  construct- 
ing the  new  pile  to  introduce  one  or  two  layers  of  chips  or  bark,  for 


88 


THE    METALLURGY 


a  reducing  effect  upon  impurities  like  arsenic,  and  for  producing  a 
more  uniform  heat  throughout  the  pile.  Finally,  after  this  burning, 
a  large  portion  suitable  for  use  can  be  sorted  out  and  the  part  still 
incompletely  burned  can  go  to  the  next  heap. 

Stall-roasting. — In  stall-roasting,  instead  of  roasting  the  ore  in 
an  exposed  heap,  it  is  enclosed  within  walls  that  form  stalls,  and 
contain  flues  for  the  admission  of  air  to  the  ore,  and  for  the  escape 
of  smoke  to  a  main  flue  and  tall  stack.  The  strong  draft  thus  creat- 
ed shortens  the  period  of  burning,  and  also  removes  the  noxious 
fumes  to  the  upper  air.  Each  stall  consists  of  a  paved  area, 
as  shown  in  Fig.  30,  6%  by  8  ft.  and  6  ft.  high,  open  at  the  front, 


CROSS  SECTION  THROUGH  A.B. 
SCALE  X  IN.  =  1  FOOT 


a  a.— Draught  holes  connecting 
with  Flue  in  sidetcalls. 
b.b.—Flue  holes  into  Main 
Culvert. 


Fig.    29.      ROASTING  STALLS  FOR  LUMP  ORE    (SECTIONAL  ELEVATION). 


to  retain  the  contents,  which  is  closed  by  a  temporary  wall,  loosely 
built  at  each  charging. 

Fig.  29  and  30  represent,  in  sectional  elevation  and  in  plan,  a 
battery  of  stalls  for  roasting  ore.  Fifty-six  such  stalls  roast  100 
tons  of  raw  ore  daily.  Each  stall  holds  20  tons  and  requires  10  days 
to  burn  and  remove  the  contents,  which,  with  a  10%  allowance  of 
time  for  repairs,  gives  an  output  of  two  tons  per  day  each.  The 
stalls  are  built  of  rough  masonry,  or  of  slag-blocks  laid  in  clay  mor- 
tar that  are  cast  at  the  works.  The  stack  is  3Vi>  ft.  inside  diameter, 
and  75  ft.  high. 

To  charge  a  stall,  the  floor  is  prepared  with  large  irregular  pieces 
to  form  a  rough  flue  or  passage  from  front  to  back  and  two  trans- 
verse ones,  by  which  air  can  enter  at  the  bottom.  In  the  passages 
kindling  wood  is  laid.  The  remainder  of  the  floor  is  covered  with 
a  thin  layer  of  long  thin  sticks  of  wood  split  from  logs  and  poles 
(See  Fig.  30).  The  stall  is  now  filled  with  coarse  ore  and  ragging 


OF    THE    COMMON    METALS. 


89 


(1  to  y±-m.  size),  distributed  through  the  mass.  While  partly  filled, 
single  small  sticks  of  wood  are  placed  at  the  back  and  sides  as  well 
as  occasional  sticks  at  the  front.  The  front  wall  is  also  carried  up 
with  the  larger  pieces  of  ore.  The  filling  completed,  a  single  car- 
load of  ragging  is  added  above  the  ore,  then  a  3-in.  layer  of  shav- 


ings, bark,  and  chips,  then  a  layer  of  fine  ore  that  can  be  roasted  with 
care,  and  finally  a  coating  of  well-roasted  ore.  A  sheet-iron  cover 
on  this,  luted  round  the  edges  to  the  walls,  is  of  great  benefit.  The 
air  enters  at  numerous  places.  It  enters  beneath  the  ore  by  the  rough 
channels,  above  described,  through  the  interstices  of  the  temporary 


90  THE    METALLURGY 

front  wall,  and  through  openings,  a  Fig.  29,  in  the  side  walls.  The 
smoke  leaves  by  the  openings  b  into  the  main  culvert. 

The  wood  of  the  stall  having  been  ignited  at  the  front  near  the 
bottom,  the  roasting  proceeds  rapidly,  and,  by  the  end  of  the  fourth 
day,  the  heap  is  burning  throughout.  Successful  burning  is  indi- 
cated by  the  swelling  of  the  contents  and  the  rising  of  the  surface, 
sometimes  to  the  extent  of  a  foot.  Because  of  this  swelling,  the  front 
or  temporary  wall  should  be  braced  with  wooden  braces  to  oppose 
the  outward  thrust.  When,  on  the  other  hand,  the  burning  pro- 
ceeds too  rapidly,  no  such  swelling  occurs,  but  instead,  the  surface 
subsides.  Subsidence  is  due  to  the  melting  of  ore  owing  to  the  great 
heat,  and  is  an  indication  that  roasting  is  imperfect.  The  heat  can, 
however,  be  regulated  by  the  use  of  fine  ore  to  stop  cracks,  and  by 
closing  the  passages  to  the  draft  openings  a. 

If  the  ore  were  left  to  burn  and  cool  slowly,  it  would  require  15 
days  to  do  this.  In  order  to  hasten  matters,  the  front  portion  of  the 
ore  as  it  cools  may  be  removed,  taking  care  not  to  penetrate  beyond 
the  cold  portion.  Beginning  at  about  the  fourth  day  it  is  possible 
to  take  ore  away,  so  that  in  seven  or  eight  days  the  stall  is  again 
empty. 

The  ore  is  brought  to  the  stalls  by  a  track  above  the  culvert  or 
mid-flue.  While  a  stall  is  being  filled,  a  turn-plate,  as  shown  in  Fig. 
30,  is  laid  down  with  a  branch-track,  so  that  the  ore  can  be  conveyed 
above  the  stall  and  dumped  into  it.  Tracks  are  also  provided  at  the 
floor-level  of  the  stall,  and  by  these  the  roasted  ore  is  conveyed  to 
the  blast-furnaces. 

Cost  of  heap-roasting. — We  find  the  cost  of  roasting  at  Duck- 
town,  Tennessee,  to  be  42c.  per  ton.  Peters  gives  as  an  average  cost 
for  fuel,  labor,  and  supplies,  48. 5c.  per  ton,  with  common  labor  com- 
puted at  $1.50  per  day.  Heap-roasting  often  can  be  done  by  con- 
tract to  advantage.  At  the  United  Verde,  Jerome,  Arizona,  75c.  per 
ton  was  the  contract  price. 

Cost  of  roasting-stalls. — Peters  gives  for  the  cost  of  building  56 
stalls  a  total  of  $3303.80,  or  about  $60  per  stall.  To  the  total  should 
be  added  $400  for  the  cost  of  the  stack. 

Cost  of  roasting  in  stalls. — This  may  be  estimated  at  50c.  per  ton, 
with  common  labor  at  $1.50  per  10-hour  day. 

Relative  advantage  of  heap  and  of  stall-roasting. — Heap-roasting 
has  the  advantage  that  it  requires  only  the  necessary  site,  and  needs 
no  investment  for  plant.  The  method  is  a  simple  one  and  the  result 
is  satisfactory.  On  a  small  scale,  primitive  methods  of  handling  ma- 
terials are  sufficient,  but  for  a  large  scale  we  must  not  forget  the  cost 


OF    THE    COMMON    METALS.  91 

of  grading,  of  trestles,  of  tracks,  etc.  Stall-roasting  saves  much 
time,  requiring  10  days  as  against  70  days  for  heap-roasting.  In 
large  plants,  where  from  10,000  to  50,000  tons  are  in  process  of  treat- 
ment, heap-roasting  may  cause  the  locking  up  of  several  hundred 
thousand  dollars  in  the  heaps.  By  reducing  this  to  one-seventh  by 
stall-roasting,  an  important  saving  is  effected.  In  stall-roasting  the 
stack  removes  the  fume,  and  the  entire  contents  of  the  stall  becomes 
roasted,  including  the  fine,  which  is  more  thoroughly  roasted  than  in 
heaps.  The  elimination  of  sulphur  is  perhaps  a  little  less  thorough 
in  stalls  than  in  heaps.  In  stalls,  rain  and  snow  have  little  effect  on 
the  process,  and  in  a  moist  climate  the  consequent  leaching  causes 
no  trouble.  For  stall-roasting  one-fifth  cord  of  wood  per  stall  is 
enough  for  charging,  or  1%  the  weight  of  the  ore  roasted,  while  in 
heap-roasting  average  practice  calls  for  2%  per  cent. 

18.     ROASTING  OF  ORES  IN  PULVERIZED  CONDITION. 

General. — This  work  is  performed  in  furnaces,  generally  of  the 
reverberatory  type,  the  ore  being  exposed  upon  a  hearth  to  the  ac- 
tion of  the  flame  and  air.  The  reactions  by  which  the  ore  is  roasted 
are  given  in  section  13,  on  'Chemistry  of  roasting'.  The  ore  if  not 
already  fine,  is  crushed  to  the  size  designated  under  'Crushing  for 
treatment  in  roasting  furnaces'.  In  these  various  furnaces  advan- 
tage is  taken  of  the  heat  developed  by  the  oxidation  of  the  sulphides. 
If  the  percentage  of  sulphur  is  high,  this  is  often  enough  to  supply 
the  required  heat  (after  combustion  has  once  started)  without  the 
aid  of  extraneous  fuel.  Thus,  in  the  McDougall  roaster,  after  the 
furnace  has  been  heated,  an  ore  containing  25  to  30%  sulphur  con- 
tinues to  roast  by  its  own  heat. 

The  various  mechanical  furnaces  roast  ore  cheaply,  but  for  ores 
containing  lead,  which  agglomerate,  roasting-furnaces  that  are 
mechanically  stirred  do  not  give  such  satisfaction  as  hand-rever- 
beratory  roasters.  With  a  slight  accession  of  heat  above  the  normal, 
caused  by  the  lack  of  care  in  firing,  the  ore  is  liable  to  agglomerate, 
and  eventually  to  stick  to  and  collect  upon  the  hearth.  In  the  hand- 
reverberatory  roaster  the  hearth  is  accessible,  and  when  this  oc- 
curs the  furnace  can  be  used  until  the  matter  becomes  serious,  then 
the  accumulation  can  be  removed  by  'cutter-bars'.  On  the  other 
hand,  such  an  accumulation  soon  stops  the  movement  of  a  mechanical 
roaster.  To  remove  it  a  flat  bar  of  iron  may  be  attached  to  one  of 
the  rabble-arms  in  the  place  of  a  rabble-blade.  This  is  made  stout 
enough  to  plow  up  the  accretions  and  by  setting  it  in  different  posi- 
tions on  the  arm,  the  hearth  is  finally  cleared.  This  device  has  not, 


92  THE    METALLURGY 

however,  proved  to  be  altogether  successful.  The  hand-reverbera- 
tory  works  well  on  ores  that  need  a  high  finishing  heat  to  break  up 
the  sulphates.  Zinc  ores  are  of  this  kind.  Such  temperatures  are 
destructive  to  any  kind  of  mechanical  iron  stirrer. 

We  may  classify  roasting-furnaces  into:  (A)  hand,  and  (B) 
mechanical-roasters. 

19.     HAND-OPERATED  ROASTERS. 

The  long-hearth  reverberatory-roaster  or  calciner. — In  this  fur- 
nace the  charge  is  dropped  and  removed  at  intervals.  The  essential 
features  (See  Fig.  31  and  32),  are  a  floor,  or  hearth  upon 'which  lies 
the  ore  spread  over  the  entire  surface  of  large  area,  a  'fire-box'  at 
one  end,  a  space  below  the  fire-grate  called  an  'ash-pit',  and  a  wall 
saparating  the  fire-box  from  the  hearth  called  a  'fire-bridge'  (or 
simply  a  'bridge').  The  whole  furnace  is  covered  by  a  flat  arch,  or 
'roof,  against  which  the  heat  of  the  flame  is  reverberated  or  thrown 
down,  upon  the  charge.  Thus  the  flame  imparts  heat  during  its  en- 
tire passage  over  the  ore  to  the  outlet-flue  at  the  opposite  end  of  the 
hearth,  where  it  escapes  by  a  flue  to  the  stack.  These  furnaces  are 
distinguished  from  the  reverberatory  smelting-furnaces  shown  in 
Fig.  139,  by  the  relatively  small  grate-area,  and  the  flat  hearth  at 
the  level  of  the  door-sills. 

The  figures  represent,  in  sectional  plan  and  in  longitudinal  sec- 
tional elevation,  a  long-bedded  reverberatory  hand-roaster.  Disre- 
garding the  fuse-box,  Fig.  33  gives  a  good  idea  of  its  appearance. 
The  width  inside  for  convenience  in  stirring  and  moving  the  charge, 
should  be  14  ft.,  while  the  length  is  designed  according  to  the  char- 
acter of  the  ore  it  is  to  treat.  The  length  of  the  hearth  must  accord 
with  the  quantity  of  sulphide  the  ore  contains,  and  consequently  the 
heat  the  ore  developes  in  roasting.  Without  the  oxidation  of  sul- 
phur the  fire  would  not  maintain  enough  heat  to  roast  the  ore  more 
than  32  ft.  distant  from  the  fire-bridge.  The  heat-generating  power 
of  the  ore  depends  upon  the  percentage  of  sulphur  contained  but  is 
greater  when  the  sulphur  is  in  the  form  of  the  loosely-held  first 
equivalent.  An  ore  containing  only  10%  sulphur  would  be  roasted 
to  good  advantage  in  a  short  furnace.  The  length  of  the  first  hearth 
of  the  furnace  shown  in  Fig.  31  is  16  ft.  Where  15%  sulphur  is  pres- 
ent it  is  proper  to  add  another  hearth,  making  a  furnace  32  ft.  long. 
A  20%  ore  would  work  rapidly  on  a  three-hearth  furnace ;  and  ore 
containing  25%  sulphur  or  more,  a  four-hearth  furnace,  as  shown 
in  the  figure,  making  a  total  length  of  hearth  of  64  ft.,  which  is  suf- 


OF    THE    COMMON    METALS. 


93 


ficient  for  any  ordinary  ore.     Hearths  of  greater  length  have  been 
tried,  but  have  not  been  found  satisfactory. 

The  stack,  proposed  by  Peters  to  furnish  draft  for  two  roasters, 


\ . 
x.  > 
s  <. 


is  42  in.  square  inside  and  65  ft.  high.  The  cost,  together  with  that 
of  the  connecting  flues,  he  gives  as  $728,  and  the  cost  of  each  rever- 
beratory  roasting-furnace,  $2713.  '  A  two-furnace  roasting-plant 


94  THE    METALLURGY 

with  building  and  accessories  would  cost  approximately  $10,000. 

This  type  of  furnace  has  several  advantages  in  the  roasting  of 
ore.  The  operation  is  started  at  a  low  temperature  at  which  there 
is  but  little  tendency  for  ore  to  cohere  or  agglomerate.  The  pulver- 
ous  condition  causes  a  thorough  contact  with  the  air,  and  makes  it 
easy  to  rabble  the  ore.  There  is  a  saving  in  fuel,  for  this  length  of 
furnace  reduces  the  temperature  of  the  escaping  products  of  com- 
bustion to  270 °C.  There  is  thorough  stirring  and  turning,  resulting 
from  the  movement  of  the  charge  toward  the  fire-end  of  the  hearth. 
The  firing  is  uniform  and  there  is  economy  in  repairs  due  to  the  uni- 
form and  moderate  heat  of  the  furnace  at  the  rear  where  red  brick 
can  be  used.  Moreover  the  heat  near  the  bridge  can  be  readily  in- 
creased to  decompose  sulphates  and  to  agglomerate  if  fusible,  which 
improves  it  for  treatment  in  the  blast-furnace. 

Construction  of  furnace. — The  cast-iron  side-door  frames  are  set 
6  ft.  apart  and  opposite  (See  Fig.  33).  The  doors  are  made  from 
plates  of  sheet-iron  and  are  removable  by  means  of  a  'lifter'.  The 
floor  of  the  furnace  is  divided  into  hearths  or  divisions,  separated 
by  steps  with  a  2-in.  drop,  so  that  successive  charges,  kept  on  sepa- 
rate hearths,  do  not  mix  with  one  another  to  the  detriment  of  the 
roast.  Sometimes  these  steps  are  omitted,  but  in  that  case  the 
charges  must  still  be  kept  separate.  The  furnace  is  strongly  stayed  by 
'buckstaves'  which  are  tied  across  by  tie-rods  to  resist  the  expansion 
due  to  the  heating  of  the  brickwork,  and  to  take  the  thrust  of  the 
arched  roof.  Through  the  walls  of  the  fire-box,  openings  2%  by  4 
in.  are  often  left  for  the  admission  of  air  above  the  level  of  the  fire. 
The  bridge  has  a  passage  through  it  to  cool  it,  with  side  openings 
2%  by  4  in.  by  which  air  can  be  introduced  beneath  the  flame  and 
in  contact  with  the  ore.  These  openings  furnish  air,  which,  to- 
gether with  that  which"  enters  through  the  fire,  produces  an  oxidiz- 
ing atmosphere  at  the  hearth.  The  fire-box,  bridge,  and  the  first  16 
ft.  of  the  hearth  should  be  of  fire-brick.  Beyond  this,  red  brick  of 
the  better  quality  may  be  used,  both  for  the  roof  and  for  the  pave- 
ment of  the  hearth.  All  brick  must  be  laid  in  clay,  not  in  lime-mor- 
tar. A  charge  of  ore  is  kept  in  the  hopper  at  the  flue-end  of  the 
furnace  ready  to  be  dropped  upon  the  hearth  when  needed.  The 
finished  charge  is  removed  through  the  discharge-opening  into  a 
car  or  wheelbarrow  standing  below. 

Operation  of  furnace. — Assuming  that  the  furnace  is  in  regular 
operation,  the  charge  on  the  first  hearth  is  withdrawn  by  pushing 
aside  a  square  cast-iron  plate  that  covers  the  discharge-opening; 
and  with  the  aid  of  a  long-handled  paddle  and  rabble,  the  roasted 


OF    THE    COMMON    METALS. 


95 


charge  on  the  first  hearth  is  raked  into  the  tram-car  beneath.  The 
paddle  used  for  the  purpose  has  a  handle  made  of  1^4-in.  pipe  16  ft. 
long.  It  has  a  blade  6  in.  wide  by  18  in.  long.  The  rabble,  really  a 


•jams 


96  THE    METALLURGY 

hoe,  has  a  handle  of  the  same  length  and  a  blade  6  in.  wide  by  10  in. 
long.  The  hearth  being  thus  cleared,  the  charge  on  the  second 
hearth  is  transferred  to  it  by  means  of  the  paddle.  In  the  same  way 
the  content  of  the  third  hearth  is  transferred  to  the  second,  and  the 
fourth  to  the  third,  thus  leaving  the  last  hearth  empty.  The  charge 
is  dropped  in  from  the  hopper,  and  by  means  of  the  paddle,  spread 
out  on  the  fourth  hearth.  This  moving  down  of  the  charge  occurs 
every  4  hours,  so  that  if  we  charge  2  tons,  we  roast  12  tons  daily. 
From  the  12  tons  daily  charged  into  the  furnace  we  may  obtain  10.2 
tons  of  roasted  ore,  the  difference  in  weight  being  due  to  the  loss  of 
sulphur,  etc.  Besides  the  movement  necessary  to  advance  the  ore 
through  the  furnace  the  charge  must  be  stirred  or  raked  at  least 
every  20  minutes.  The  proper  fuel  for  the  furnace  is  a  free-burning 
semi-bituminous  coal,  which  should  be  burned  in  a  shallow  bed,  6  to 
8  in.  thick  upon  the  grate,  and  it  should  be  added  every  15  minutes. 
A  roaster  consumes  5000  to  6000  Ib.  coal  per  24  hours.  With  an  out- 
put of  12  tons  daily  this  is  a  consumption  of  21  to  25%  of  the  ore 
charged.  The  cost  of  roasting  a  copper  sulphide  ore  in  a  long- 
bedded  reverberatory  furnace  is  $1.81  per  ton  ore  charged.  If  the 
ore  contains  lead  (which  makes  it  slow  and  more  difficult  to  roast), 
the  cost  may  rise  to  $2.25  per  ton. 

Reverberatory  furnace  with  fuse-box. — This  is  an  ordinary  long- 
bedded  roaster  to  which  has  been  added  a  slagging-hearth  or  fuse- 
box.  Fig.  33  shows  in  perspective  such  a  furnace,  75  ft.  in  total 
length.  Its  roasting-hearth  is  57  ft.  long  by  15  ft.  wide  inside.  Next 
comes  the  slagging-hearth  26  in.  lowrer,  inside  dimensions  11  by  13 
ft.  Separated  from  the  slagging-hearth  or  fuse-box  by  the  fire- 
bridge is  the  fire-box  with  grate-dimensions  of  8  ft.  by  2  ft.  10  in. 
or  23  sq.  ft.  area.  In  the  fuse-box  a  high  temperature,  sufficient  for 
melting  or  slagging  the  roasted  ore,  is  produced,  so  that  the  fire- 
bridge must  be  furnished  with  a  water-jacket,  or  coil  of  water-cooled 
pipes  inserted  within  the  brick-work  of  the  bridge,  to  protect  it  from 
the  corrosive  or  scouring  action  of  the  slag. 

Ore,  in  the  hopper  at  the  fuel-end,  is  dropped  upon  the  bed  of 
the  furnace,  and  roasted  as  in  the  ordinary  roaster.  When  roasted 
to  this  extent,  it  is  discharged  into  the  fuse-box  where  the  tempera- 
ture is  high  enough  to  melt  it  down.  When  melted  it  is  skimmed  or 
withdrawn  by  means  of  a  rabble.  Certain  ores  containing  blende, 
that  in  roasting  produces  sulphate  difficult  to  decompose,  are  roasted 
and  then  slagged.  Silica,  which  may  be  in  the  ore,  or  may  be  added 
in  the  fuse-box,  reacts  according  to  reaction  (13),  or  when  lead  sul- 
phate is  present  by  reaction  (14),  thus  eliminating  sulphur.  The 


OF    THIS    COMMON    METALS.  97 

chief  objection  to  such  treatment  is  that  there  is  loss  of  silver  and 
lead  as  mentioned  under  'Chemistry  of  roasting'. 

20.     MECHANICAL  ROASTERS. 

These  may  be  divided  into : 

(1)  Revolving  cylinders,  set  nearly  or  quite  horizontal,  and  re- 
volving on  the  long  axis  with 

(a)  Continuous    discharge,    as    the    White-Howell    and    the 
Argall. 

(b)  Intermittent  discharge,  as  the  Bruckner. 

(2)  Automatic  reverberatory  roasters  or  calciners  with  continu- 
ous discharge  having : 

(a)  Straight-hearths,  as  the  Brown-O 'Harra,  the  Wethey,  the 
Edwards,  the  Merton. 

(b)  Curved  or  circular-hearth,  as  the  Brown-horseshoe,  the 
Pearce-turret,  and  the  McDougall. 

Besides  these  may  be  mentioned  the  Holthoff  and  the  Raymond, 
the  hearth  of  which  itself  revolves,  and  also  the  Stetefeldt  shaft- 
furnace  in  which  the  ore  is  showered  down  a  shaft. 

The  White-Howell  furnace. — Fig.  34  is  a  longitudinal  elevation  of 
this  furnace.  It  consists  of  a  cylinder,  50  in.  inside  diameter  by  34 
ft.  long,  set  at  an  inclination  of  21/2%,  supported  on  friction-rollers 
carried  on  the  driving  shaft.  At  one  end  is  the  fire-box,  at  the  other 
a  dust-chamber  which  connects  by  a  flue  to  the  stack.  The  hotter 
end  of  the  cylinder,  near  the  fire-box,  is  of  larger  diameter,  to  per- 
mit of  its  being  lined  with  brick,  thus  leaving  the  cylinder  of  uni- 
form interior  diameter  throughout.  Projecting,  longitudinal,  fire- 
brick ledges,  set  spirally,  raise  the  ore  and  shower  it  back  through 
the  flame  as  the  cylinder  revolves,  so  as  to  roast  it  more  rapidly.  The 
unlined  part  for  the  same  reason  is  furnished  with  longitudinal,  cast- 
iron,  projecting  shelves.  Ore  is  fed  at  the  flue-end,  by  means  of  a 
screw-feed  (See  Fig.  194),  and  when  dropped  into  the  revolving 
cylinder,  travels  along,  discharging  at  the  fire-box  end.  Just  before 
it  reaches  the  fire-box  it  passes  out  from  the  cylinder  to  a  brick 
chamber  below,  and  is  withdrawn  from  that,  when  cool.  The  furn- 
ace makes  much  flue-dust.  It  is  used  chiefly  for  chloridizing-roast- 
ing,  upon  ores  containing  but  little  sulphur,  and  has  a  capacity  of 
50  tons  per  24  hours  for  low-sulphur  ores. 

The  Bruckner  roasting-furnace. — This  consists  of  a  brick-lined 
cylinder,  supported  by  and  revolving  on  four  rollers  or  carrier- 
wheels.  As  in  the  White-Howell  furnace,  there  is  a  fire-box  at  one 
end,  and  a  flue  at  the  other.  The  flame  from  the  fire-box  is.  drawn 


98 


THE  METALLURGY 


directly  through  the  end-openings  of  the  cylinder  to  the  flue.     The 
furnace  treats  the  ore  in  charges  that  require  24  to  48  hours.     The 


OF    THE    COMMON    METALS.  99 

cylinder  is  provided  with  man-holes  for  charging  and  discharging 
the  ore,  the  ore  being  first  put  into  a  double  hopper  from  which  it 
is  quickly  drawn  when  needed. 

Fig.  35  shows  the  end  and  side  of  the  furnace,  indicating  founda- 
tions in  section.  In  the  end-view  at  the  right,  the  fire-box  is  re- 
moved. The  cylinder,  8%  ft.  diam.  by  IS1/^  ft.  long,  has  end-open- 
ings 3  ft.  diam.  for  the  entrance  and  exit  of  the  flame.  It  is  driven 
by  a  worm-gear,  the  motion  being  communicated  to  two  of  the  car- 
rier-wheels on  the  common  shaft  and  transmitted  to  the  cylinder  by 
the  friction  between  the  bearing  rings  and  the  carrier-wheels.  The 
fire-box  is  a  movable  one  that  can  be  set  aside  when  it  is  desired  to 
reach  the  interior  of  the  cylinder. 

The  present  practice  is  to  revolve  the  cylinder  slowly,  say  once 
an  hour,  since  the  contents,  even  at  this  speed,  is  constantly  shift- 
ing and  representing  new  surfaces  to  the  air.  The  old  way  was  to 
revolve  it  once  in  four  minutes.  At  the  time  of  discharging,  the 
speed  should  be  at  least  one  revolution  in  2  to  4  minutes  to  discharge 
the  ore  quickly.  This  adjustment  is  made  by  throwing  in  a  clutch 
by  means  of  a  quick-working  mechanism.  All  the  man-holes  are 
opened,  and  it  takes  but  few  revolutions  to  discharge  the  cylinder. 
The  ore  falls  into  a  pit  below,  whence  it  can  be  withdrawn  to  tram- 
cars  for  use.  Charges  that  contain  lead  are  liable  to  agglomerate, 
even  with  cautious  firing,  and  to  'hang  up',  that  is,  adhere  in  a 
layer  to  the  brick-lining.  Should  this  occur,  the  movable  fire-box 
can  be  pushed  aside  and  the  layer  removed  by  long  chisel-ended 
slice-bars.  The  attachment  to  the  brick-work  is  so  slight  that,  when 
the  layer  is  cut  from  end  to  end,  even  at  one  place,  the  key  or  con- 
tinuity is  broken  and  the  mass  falls.  Thus  it  is  easily  dislodged  and 
by  further  rolling  breaks  up  and  is  ready  for  continued  roasting.  A 
moderate  cohesion  of  the  fine  particles  does  not  present  a  good  roast. 

Operation. — The  charge  of  20  tons  having  been  dropped  into  the 
cylinder  from  the  hopper,  the  man-holes  are  closed  and  vigorous 
firing  begun  to  start  the  ore  to  burning  by  its  own  oxidation,  which 
begins  at  a  barely  visible  red.  This  takes  about  six  hours,  the  nec- 
essary temperature  being  attained  first  at  the  fire-box  end  and  ex- 
tending then  to  the  flue-end.  The  charge  thus  started  burns  by  its 
own  heat  12  hours  longer  and  the  fire  meanwhile  is  withdrawn  from 
the  fire-box.  Indeed  the  reactions  are  so  vigorous  that  the  charge 
must  be  closely  watched  and  the  air-supply  regulated  lest  agglom- 
eration occur  if  the  ore  be  a  leady  one.  As  the  heat  toward  the 
end  of  this  stage  diminishes,  firing  is  resumed,  increasing  the  heat 
gradually,  6  to  8  hours  more,  to  the  finish.  In  this  way  sulphates, 


100  THE    METALLURGY 

formed  early  in  the  roast,  are  decomposed,  and  a  good  result  is  ob- 
tained. The  whole  cycle  of  operations  takes  24  to  36  hours  for  cop- 
per ores  and  48  hours  or  more  for  leady  ores.  For  copper  ores  it  is 
sufficient  to  reduce  the  quantity  of  sulphur  to  7  to  8%,  and  for  leady 
ores  to  3  to  5%.  The  charge  is  removed  by  revolving  the  furnace 
rapidly,  and  after  10  to  15  minutes  little  is  left  to  mix  with  the  next 
charge.  To  charge  the  cylinder  again,  all  but  two  man-holes  are 
closed,  and  the  cylinder  is  revolved  to  bring  these  directly  under  the 
hopper-spouts.  The  hopper-slides  are  withdrawn  and  the  ore  is 
quickly  run  into  the  cylinder  and  is  ready  for  firing  on  closing  the 
man-holes.  These  operations  of  charging  and  discharging  need  not 
take  more  than  20  minutes.  For  copper  sulphides,  roasted  in  24 
hours,  the  furnace  capacity  accordingly  is  at  least  20  tons  daily,  and 
for  leady  sulphides  roasted  in  48  hours,  10  tons. 

The  cost  of  roasting  leady  ores  is  85c.  per  ton,  and  for  copper 
sulphides  roasted  to  7  to  8%,  42c.  per  ton.  The  cost  of  one  of  these 
roasters,  installed,  may  be  estimated  at  $3000. 

The  Wethey  roasting-furnace. — Fig.  36  is  a  perspective  view,  and 
Fig.  37  a  cross-section  of  this  furnace  which  consists  of  two,  super- 
imposed, straight  hearths,  heated  by  fire-boxes  placed  at  the  sides  of 
the  upper  hearth.  The  roasting  is  done  on  the  upper  hearth ;  and 
the  lower  hearth  is  for  cooling  the  ore  after  it  has  been  roasted. 
The  upper  one  of  the  two  hearths,  each  of  which  is  121  by  12  ft:  in 
size,  is  held  between  heavy  transverse  I-beams  above  and  below,  as 
shown  in  Fig.  37,  which  tie  the  upright  buckstaves  and  support  the 
hearth.  The  roof  of  this  hearth  is  so  suspended  from  the  upper  I- 
beam  as  to  leave  slits  the  entire  length  of  the  hearth;  and  through 
these  project  the  ends  of  the  rabble-arms  which  rest  on  carriages. 
The  stirring  blades,  or  rabbles,  within  the  furnace,  are  set  diagonally 
upon  the  rabble-arms  (see  the  detail  of  such  an  arrangement  for  a 
Pearce-turret  furnace  shown  in  Fig.  42).  The  ends  of  the  rabble- 
arms  are  attached  to  endless  chains,  which  drag  them  along  the 
upper  hearth,  stirring  and  gradually  moving  the  ore  forward.  They 
return  by  the  lower  hearth  upon  which  the  ore,  from  the  finishing 
end  of  the  upper  hearth,  falls  through  an  opening.  Here  it  is  again 
moved  along,  stirred,  cooled,  and  finally  discharged  into  a  hopper, 
thence  to  be  drawn  into  two-wheeled  buggies  and  transferred  to  the 
leaching-vats.  The  endless  chains  pass  around  sheaves  at  the  ends 
of  the  furnace,  and  around  sprocket-wheels  at  the  driving  end  which 
impel  them.  The  rabble-arms  are  so  arranged  that  they  can  be  re- 
moved readily  without  disturbing  the  carriage  connections.  There 
are  two  rabble-arms,  and  these  pass  along  the  hearths  at  the  rate  of 


OF    THE    COMMON    METALSw. 


100  ft.  per  minute.  The  blades  of  one  rabble  are  set  at  the  opposite 
angle  to  those  on  the  next,  so  that  the  ore  tends  toward  neither  side. 
To  hasten  the  cooling  of  the  ore  upon  the  lower  hearth,  that  it  may 
be  ready  promptly,  for  further  treatment,  water-cooled  pipes  are 


Fig'.   ?,6.      VIEW  OF  WETHEY   ROASTING   FURNACE. 


•*          *          # 


Fig.   37.      CROSS-SECTION  OF  WETHEY   ROASTING   FURNACE. 

laid  the  length  of  the  hearth  in  grooves  in  the  brick  pavement  flush 
with  the  top  surface.  The  ore  is  regularly  fed  to  the  furnace  from 
the  hoppers  by  automatic  feeders  at  the  driving  end  (shown  at  the 
right  in  the  perspective  view,  Fig.  36),  and  is  heated  by  two  sets  of 
fire-boxes,  one  of  which  is  at  the  feed  end,  the  other  half  way  along 


THE    METALLURGY 


the  hearth.  The  flame  from  the  fire-boxes  descends  to  the  hearth  by 
a  flue  crossing  the  roof  (See  Fig.  36),  then  moves  horizontally  to  the 
left,  where  by  a  flue  at  the  end  it  escapes  to  the  stack.  Thus  the  ore 
and  flame  move  in  the  same  direction  over  the  roasting-hearth.  At 
each  end  of  this  hearth  flap-doors  of  sheet-iron  are  hinged.  These 
hang  by  the  top  edge  of  the  sheet,  and  close  the  ends  of  the  furnace 
at  all  times  except  when  the  rabbles  enter  or  pass  out  of  the  furnace. 
The  doors  are  lifted  by  the  rabble,  but  drop  again  as  soon  as  it  passes. 
It  might  be  thought  that  the  slits,  each  120  ft.  long,  would  admit  too 
much  air  and  injure  the  draft,  but  the  furnace  is  found  to  work  well 


Fig.    38.      PLAN  AND  ELEVATION  OF  EDWARDS  ROASTING    FURNACE. 

notwithstanding  this  feature.  The  rabbles  are  in  the  open  air  more 
than  half  the  time,  and  have  a  chance  to  cool  after  each  passage 
through  the  furnace.  We  estimate  the  capacity  of  this  furnace, 
roasting  ore  containing  Vi2  to  3y^%  sulphur  that  needs  13  to  15  sq. 
ft.  hearth-area,  at  100  tons  per  24  hours. 

The  Edwards  roasting-furnace. — This  is  a  single-hearth  rever- 
beratory  furnace  with  hearth  dimensions  57  ft.  long  by  6  ft.  wide. 
Fig.  38,  in  plan,  shows  a  portion  at  the  fire-box  end,  the  feeding 
mechanism  and  the  cooling  floor  in  section.  The  elevation  shows 
the  side,  constructed  like  a  plate-iron  beam,  the  stirring  mechanism 
and  the  conveyor  for  transferring  the  roasted  ore  to  the  cooling-pit. 
F1g  )9  is  a  transverse  section  of  tho  hearth  showing  the  details  of 


OF    TIIK    COMMON    METALS. 


103 


the  stirring  mechanism.  The  slope  of  the  furnace  can  be  changed  a 
little  by  tilting.  This  regulates  the  rate  of  travel  of  the  ore  through 
the  furnace ;  but  for  a  given  kind  of  ore,  this  slope,  once  determined,- 
is  not  again  changed.  The  furnace  has  a  slope  of  2  in.  per  foot  to- 
ward the  discharge  or  fire-box  end.  The  stirring  and  propulsion  of 
the  charge  is  affected  by  means  of  rabbles  fixed  to  vertical  shafts,  as 
shown  in  the  elevation  of  Fig.  39,  and  in  the  plan  of  Fig.  38.  The 
rabbles  at  the  fire-box  end  are  water-cooled,  and  this  is  found  especi- 
ally necessary  where  a  high  finishing  heat  is  needed.  The  blades  or 
plows  of  the  rabbles  can  be  easily  replaced  through  the  doors  adja- 
cent to  them.  The  figure  indicates  the  hearth  as  broken  away,  at 
the  discharge  end,  to  show  two  of  the  rabbles  in  plan.  The  last 


Fig.    39.      CROSS-SECTION   OF   EDWARDS  ROASTING   FURNACE. 


rabble  sweeps  the  roasted  ore  into  the  discharge  shoot,  and  the  push- 
conveyor  then  moves  the  ore  to  the  cooling-pit.  The  bottom  of  the 
conveyor-trough  is  furnished  with  slides,  by  means  of  which  the  ore 
can  be  dropped  at  any  desired  point  on  the  cooling-floor.  The  ore 
is  fed  to  the  furnace  from  the  feed-hopper,  by  an  endless-screw  con- 
veyor which  discharges  into  a  feed-opening  in  the  roof  of  the  fur- 
nace. The  smoke  is  carried  off  by  a  flue.  The  furnace  takes  1  hp.  to 
operate,  and  has  a  daily  capacity  of  25  tons  on  sulphide  ore  of  30  to 
35%  sulphur.  The  roasted  ore  contains  3  to  8%  of  sulphur.  The 
moving  parts  are  durable,  and  the  furnace  has  proved  efficient  in 
practice.  Large  installations,  of  the  duplex  type  with  a  double  in- 
stead of  a  single  row  of  rabbles,  and  of  hearth-dimensions  120  by  12 
ft.,  have  been  built  for  a  daily  capacity  of  60  tons.  These  furnaces 
do  not  have  the  tilting  hearth. 


104 


TTTIO    METALLURGY 


The  Pearce-turret  roasting-furnace. — The  word  'turret'  has  ref- 
erence to  the  circular  form  of  the  furnace.  Three  types  have  been 
developed,  namely,  the  one-deck,  the  two-deck,  and  the  six-deck  or 


Fig.    40.      TWO-DECK    PEARCE-TURRET    ROASTING    FURNACE    (PL.AN). 

six-hearth  furnace.  The  greater  the  number  the  hearths,  the  more 
economical  the  furnace  is  as  regards  fuel  and  output.  On  the  other 
hand,  the  multiplicity  of  hearths  makes  it  more  complicated,  and 


106  TIIK    METALLURGY 

causes  more  flue-dust  by  the  repeated  dropping  of  the  ore  from 
hearth  to  hearth. 

We  shall  now  consider  the  two-deck  type  shown  in  Fig.  40  and 
41.  Fig  40  is  a  plan,  with  opposite  portions  broken  away  to  expose 
the  hoppers  and  the  rabble.  Fig.  41  is  a  central  sectional  elevation 
through  a  fire-box  and  the  hearths.  To  understand  these  views  will 
require  careful  study.  There  are  two  superimposed  annular  hearths 
with  space  between  for  the  15-in.  I-beam  supporting  the  upper  hearth 
and  a  flue-chamber  below.  Referring  to  the  plan,  the  ore-hopper, 
partly  broken  away,  is  shown  at  the  near' side.  Next,  at  the  left,  is 
the  outlet-flue  with  its  barred  top,  for  the  escaping  gases ;  and  at  the 
right,  where  the  furnace  is  broken  down  to  the  lower  hearth,  is 
shown  the  discharge-hopper.  The  ore  is  fed  continuously  from  the 
ore-hopper  (See  Fig.  41),  through  the  roof  to  the  upper  hearth  until, 
having  made  its  circuit,  it  falls  through  a  transverse  slit  to  the  lower 
hearth.  Here,  as  before,  it  is  stirred  by  rabbles  and  moved  forward 
until,  having  again  made  the  circuit,  it  is  received  into  the  discharge- 
hopper  of  the  lower  hearth. 

There  are  two  rabble-arms  for  each  hearth.  Of  these,  the  upper 
two  are  shown  in  the  plan.  The  lower  two  are  at  right  angles  to  the 
upper  ones  and  are  shown  in  the  elevation.  Each  pair  of  rabble- 
arms  is  screwed  into  a  central  hub,  that  revolves  on  a  fixed  central 
column.  A  gear-wheel  on  the  rabble-arms,  in  connection  with  a 
driving  mechanism,  moves  the  rabble  at  the  rate  of  about  one  revolu- 
tion in  two  minutes.  As  will  be  noticed  in  the  plan,  Fig.  40,  and 
in  the  detail,  Fig.  42,  the  rabble-blades  are  set  at  an  angle  to  their 
line  of  travel.  Nearly  touching  the  hearth  they  pass  through  the 
ore,  stirring  it  and  moving  it  slightly  forward  in  pushing  it  aside. 
The  blades  of  the  opposite  rabble-arm  incline  in  the  opposite  direc- 
tion and  stir  the  ridges  thus  formed.  Thus  every  minute  fresh  sur- 
face is  exposed  to  the  action  of  the  fire.  An  objection  has  been 
urged  against  this,  as  against  all  the  mechanically  stirred  roasters 
(except  the  Edwards),  that  a  part  of  the  imperfectly  roasted  ore, 
adhering  to  the  blades,  creeps  forward  and  mixes  with  the  advanced 
portions  of  the  charge,  raising  the  percentage  of  sulphur  in  a  way 
that  need  never  occur  in  hand-roasting.  Air  from  a  fan-blower  is 
forced  into  the  central  hollow  column,  thence  out  through  the  rab- 
ble-arms, thus  cooling  them,  finally  through  nipples  screwed  into 
the  underside  of  the  arm  between  the  blades,  as  shown  in  Fig.  42. 
In  addition  to  this  forced  air  from  the  fan,  air  enters  through  open- 
ings at  the  sides  to  oxidize  the  ore. 

The  fire-boxes,  one  of  which  is  well  shown  in  Fig.  41,  are  sup- 


OF    THE    COMMON    METALS. 


107 


plied  by  air  from  the  fan-blower  under  a  slight  pressure.  By  means 
of  the  under-grate  blast  the  flame  is  caused  to  enter  the  furnace 
under  a  slight  pressure,  and  thus  oppose  the  inward  draft  of  air 
through  the  side-doors.  Suction  increases  as  we  approach  the  exit- 
flue,  as  indeed  it  must  in  order  to  take  away  the  products  of  com- 
bustion. The  fire-box  is  furnished  with  a  step-grate  suitable  for 
burning  slack  coal.  The  ash-pit  is  closed  by  tight  sheet-iron  doors, 
so  that  the  under-grate  blast  can  be  sustained.  Occasionally  these 
are  opened  so  that  the  firemen  can  clean  the  grate.  The  ashes  and 
clinkers  are  dislodged  from  between  the  step-grates,  falling  into  the 


rta.  9. 


,'    tNVCPTtO  JftMt. 

Fig.    42.      DETAIL,  OF   RABBLES. 

ash-pit  arid  being  thence  removed.  The  fire-box  is  supplied  with  coal 
by  means  of  a  hopper  kept  constantly  full.  To  prevent  an  intense 
action  of  the  heat  where  the  flame  would  naturally  strike  the  charge, 
a  curtain-arch,  lower,  and  having  less  curvature,  as  seen  in  Fig.  41, 
deflects  and  distributes  the  flame.  There  are  three  fire-boxes,  all 
of  which  are  upon  the  upper  hearth ;  and  the  movement  of  the  flame 
is  in  a  direction  contrary  to  that  of  the  rabbles  and  the  ore.  The 
flame  from  the  fire-box  nearest  the  feed  moving  toward  the  outlet- 
flue  heats  the  ore  and  sets  it  afire.  The  second  and  third  fire-boxes 
raise  temperature  of  the  ore  to  the  full  heat ;  but  the  final  action  on 
the  lower  hearth  is- performed  by  the  heat  from  the  oxidation  of  the 
ore.  The  outlet-flue  leads  from  the  roof  of  the  upper  hearth,  down 
the  side  of  the  furnace,  to  the  dust-chamber  below.  An  under- 


108  THE    METALLURGY 

ground-flue  from  the  dust-chamber  leads  to  the  chimney.  The 
hearths  are  bound  by  heavy  bands.  The  upper  hearth  and  the 
mechanism  of  the  furnaces  are  sustained  by  the  I-beam  construction 
that  serves  also  to  provide  for  the  slit  through  the  inner  wall  of  the 
hearth  through  which  the  rabble-arms  must  enter.  The  slit  is  cov- 
ered by  a  sheet-metal  band,  which  is  attached  to  and  revolves  with 
the  rabble-arms.  The  question  might  be  asked,  whether  the  flame 
from  the  third  fire-box  would  not  go  directly  to  the  outlet-flue  in- 
stead of  following  round  the  hearth.  To  prevent  this,  the  upper 
hearth  is  broken  or  interrupted  for  a  space  of  5  feet  along  the  hearth, 
the  rabbles  being  visible  for  inspection  as  they  pass  through  this 
open  section.  The  ends  of  the  hearth  are  closed  by  swinging  sheet- 
iron  doors,  like  those  of  the  Wethey  furnace.  As  the  rabble  enters 
the  open  space  the  doors  are  lifted  by  the  rabble-arm,  and  drop  by 
gravity  when  it  passes  on.  Though  complicated  in  design,  this  is  a 
furnace  that  has  worked  well  in  practice.  It  is  33  ft.  interior  diame- 
ter and  has  a  6  ft.  hearth,  this  giving  for  the  hearths  an  area  of 
1020  sq.  ft.  Its  daily  capacity  is  35  tons  of  ore  of  35%  sulphur, 
which  it  roasts  to  6  to  1%  sulphur  with  a  consumption  of  9.1%  fuel. 
The  labor  needed  is  but  little  more  for  a  double-deck  furnace  than 
for  the  single  one,  while  the  fuel-cost  per  ton  of  ore  is  reduced  one- 
half.  Thus  a  single-deck  furnace  consumes  18%  of  fuel.  More  flue- 
dust  is  made  in  the  double-deck  furnace.  To  operate  the  two-deck 
furnace  requires  3  hp.,  and  the  cost  of  roasting  is  98c.  per  ton.  Such 
a  roaster  costs  $8000  installed.  In  a  multiple,  or  six-deck  furnace, 
the  fuel  has  been  reduced  to  1.4%  of  the  ore  tonnage,  but  the  flue- 
dust  is  increased  to  4  per  cent. 

The  McDougall  roasting-furnace. — There  are  several  kinds  of 
furnaces  of  this  type.  Among  these  are  the  Herreshoff,  the  Wedge, 
and  the  Evans-Klepetko.  The  latter,  as  manufactured  by  the  Allis- 
Chalmers  Co.,  is  shown  in  sectional  elevation  in  Fig.  43.  It  is  a 
vertical,  cylindrical  furnace,  16  ft.  diam.,  with  six  arched  hearths, 
over  which  travel  rabbles  which  stir  and  move  the  ore  gradually 
toward  the  central  drop-opening  through  the  floor  of  each  hearth, 
situated  alternately  at  the  center  and  at  the  periphery.  A  centra  I 
shaft  is  provided,  carrying  six  radial  rabble-arms  (three  of  these  arc 
hidden  by  the  shaft  in  the  illustration),  provided  with  nibble-blades 
set  at  an  angle  on  the  arm  as  shown  in  Fig.  42. 

The  rabble-blades  on  the  even-numbered  hearths  are  so  set  as  to 
push  the  ore  toward  the  periphery ;  the  odd-numbered  ones  toward 
the  center.  The  ore,  fed  continuously  into  the  furnace  from  a  cylin- 
drical hopper  shown  above  and  at  the  right,  Fig.  43,  drops  upon  the 


OF    THE    COMMON    METALS. 


109 


upper  hearth  near  its  outer  edge.  The  rabble-blades  of  that  hearth 
stir  and  move  the  ore  gradually  toward  the  central  drop-opening 
where  it  falls  to  hearth  No.  2.  The  rabbles  of  this  hearth  again  stir 
and  move  it  to  the  outer  drop-openings,  through  which  it  falls  to 
hearth  No.  3.  The  ore  advances  by  this  means  until  it  reaches  the 
lower  hearth,  where  an  opening  at  the  periphery  gives  it  exit  to  a 


Fig.    43.      McDOUGALL  ROASTING   FURNACE    (ELEVATION). 

receiving-hopper,  shown  beneath  the  hearth,  from  which  it  is  drawn 
into  a  car  as  is  required. 

A  high-sulphide  ore  roasts  by  its  own  heat  when  the  furnace  is 
in  full  operation.  The  ore  fills  the  hearth  to  the  level  of  the  blades, 
and  is  spread  out  evenly  by  them.  On  the  upper  hearth,  as  the  ore 
moves  toward  the  central  opening,  it  becomes  dry  and  hot,  and  when 
dropped  upon  hearth  No.  2,  begins  roasting.  On  hearth  No.  3,  the 


110  THE    METALLURGY 

ore  roasts  freely  emitting  sparks  and  forming  sulphates.  On  hearth 
No.  4  no  sparks  are  seen,  and  the  ore  has  attained  its  highest  tem- 
perature. On  hearth  No.  5  the  ore  looks  less  bright;  and  on  No.  6, 
especially  at  the  discharge,  it  has  become  cooler. 

The  air  for  oxidation  is  admitted  by  side-doors,  mostly  those  of 
the  lower  hearths.  The  gas,  and  dust,  passing  up  through  the  drop- 
openings,  are  drawn  through  the  horizontal  main  flue.  In  starting, 
the  furnace  is  heated  to  the  kindling  temperature  of  the  ore  which, 
if  rich  in  sulphur,  burns  by  its  own  heat,  without  the  aid  of  fuel.  If 
the  sulphur  content  is  low,  additional  heat  is  supplied  by  one  or 
more  external  fire-places,  near  the  bottom  of  the  furnace. 

To  protect  the  rabble-arms  from  the  intense  heat  they,  and  like- 
wise the  central  shaft,  are  water-cooled.  The  cooling-water  is  forced 
down  the  9-in.  hollow,  central  shaft  in  a  3-in.  pipe  to  a  point  near 
the  bottom,  and  out  to  the  ends  of  the  arms  in  1-in.  pipes.  It  then 
returns  up  the  annular  space  between  the  3-in.  pipe  and  the  hollow 
shaft,  and  discharges  at  the  top  through  two  spouts  into  a  launder. 
The  furnace  is  IS1^  ft.  high  by  16  ft.  diam.,  and  has  a  total  hearth- 
area  of  nearly  1000  sq.  ft.  The  structure  is  supported  on  columns  to 
give  room  below  for  the  hopper  and  the  car  into  which  the  roasted  ore 
is  discharged.  The  shell  made  of  %-in.  plate-steel,  is  lined  with  9  in. 
of  brick-work.  The  rabble-arms  consume  I1/!'  to  2  hp.,  and  make  one 
revolution  in  1%  minutes. 

A  furnace  treats,  in  24  hours,  40  tons  of  sulphide  ore  of  35% 
sulphur,  reducing  it  to  7  per  cent.  About  4%  flue-dust  is  made ;  and 
the  ore  itself  contains  more  ferric  oxide,  and  is  lighter  and  more 
porous  than  if  treated  in  a  hand-reverberatory  roaster.  The  cost  of 
roasting  such  ore  is  approximately  35c.  per  ton,  which  is  the  lowest 
figure  thus  far  known  for  any  furnace.  The  compact  form  of  the 
furnace  reduces  radiation  to  a  minimum  and  enables  roasting  with 
little  or  no  fuel.  Taking  capacity  into  consideration,  the  furnace  is 
one  of  moderate  price,  and  one  that  costs  little  to  keep  in  repair. 

Of  the  two  revolving-hearth  furnaces  the  Holthoff  and  the  Ray- 
mond, the  latter  has  some  popularity  for  the  preliminary  roasting  of 
ores  for  lime  or  pot-roasting,  the  powdered  ore  being  showered  down 
a  vertical  shaft  or  tower  and  coming  in  contact  with  an  upward 
flame  from  a  fire-box.  An  objection  to  its  use  is  the  flue-dust  that 
is  made. 

21.     ROASTING  OF  MATTE. 

Copper-bearing  matte  is  not  difficult  to  roast,  but  must  be  crushed 
at  least  to  4-mesh  size. 


OF    THE    COMMON    METALS.  Ill 

The  term  'roasting'  is  applied  also  to  a  method  of  treating  cop- 
per matte  in  a  reverberatory  furnace  in  large  pieces,  upon  which 
an  oxidizing  flame  is  allowed  to  play.  Such  masses  slowly  melt  and 
are  acted  on  by  the  air,  whereby  a  part  of  the  material  becomes 
oxidized  or  roasted  sufficiently  for  the  next  operation.  As  com- 
pared with  ordinary  roasting  this  is  slow,  and  the  method  is  one 
but  little  used.  Lead-bearing  matte  from  the  silver-lead  blast-furn- 
ace, to  be  roasted  in  a  reverberatory  furnace  of  the  kind  shown  in 
Fig.  32,  needs  a  different  treatment  from  that  given  to  ore.  This 
kind  of  matte  contains  but  20%  sulphur,  and  does  not  take  fire  like 
pyrite  ore,  but  must  have  a  high  finishing  heat  to  expel  the  sulphur. 
Such  matte  is  considered  well  roasted  when  it  contains  4%  sulphur. 
Ores  low  in  lead  can  easily  be  roasted  to  2  to  3%  sulphur,  while 
galena,  when  roasted,  still  contains  5  to  6%  when  drawn  from  the 
furnace.  Like  matte,  galena  starts  burning  slowly,  and  must  be 
roasted  slowly,  for  rapid  heating  at  once  causes  it  to  sinter  and  thus 
stops  further  roasting.  Typical  leady  matte  contains  metals  and 
sulphur  as  follows : 

Raw,  low-grade.     Eoasted,  low-grade.     Raw,  shipping. 
Per  cent.  Per  cent.  Per  cent. 

Pb    10.66  10.49  9.06 

Cu    4.62  4.12  42.30 

Fe    53.11  52.41  20.00 

S  26.87  6.13  17.89 


95.26  73.15  89.25 

The  roasted,  low-grade  matte  tabulated  above  contains  23% 
oxygen.  This  explains  why  it  does  not  lose  weight  in  roasting. 
Pyrite  ores  of  20  to  30%  sulphur,  on  the  contrary,  easily  lose  15%  in 
weight. 

22.     LOSSES  IN  ROASTING. 

Such  loss  depends  upon  the  extreme  to  which  the  roasting  is  car- 
ried as  well  as  upon  the  nature  of  the  ore.  When  ore  is  so  roasted 
that  it  is  not  sintered  at  the  final  high  temperature,  the  lead  lost 
averages  2.5%  but  no  loss  of  silver  occurs.  When  the  temperature 
is  carried  higher,  and  the  ore  is  agglomerated,  the  loss  is  slightly 
higher.  When  fused  it  may  reach  15  to  20%  of  the  lead  and  2  to 
5%  of  the  silver.  Of  the  gold  little  is  lost  in  oxidizing  roasting. 

23.  CAPACITY  OF  FURNACES  AND  COST  OF  ROASTING. 

These  depend  upon  the  surface  exposed  to  the  oxidizing  influ- 


112  THE    METALLURGY 

ences  and  upon  the  quantity  of  sulphur  contained  in  the  ore.  Sili- 
cious  ore,  containing  !/>  to  3l/2%  sulphur,  requires  13  to  15  sq.  ft.  of 
hearth-area  per  ton  ore  roasted  per  24  hours.  Matte  containing  20 
to  25%  sulphur,  when  it  is  required  to  reduce  the  sulphur  content 
to  4%,  needs  45  sq.  ft.  hearth-area;  copper  sulphide  ore,  roasted  to 
1%  in  preparation  for  smelting,  requires  33  to  35  sq.  ft.  For  roast- 
ing iron-sulphide  concentrate,  which  carries  35  to  45%  sulphur, 
down  to  3  to  10%  sulphur,  55  to  60  sq.  ft.  hearth-area  is  needed. 

Ore-roasting  in  heaps,  at  Jerome,  Arizona,  costs  80c.  per  ton,  in- 
cluding general  expense.  Ore-roasting  in  stalls  costs  50c.  per  ton. 
For  reverberatory  roasting,  in  long,  hand-rabbled  furnaces  the  low- 
est price  attainable  on  copper  ores  was  $1.50,  with  an  average  of 
$1.81  per  ton.  For  roasting  lead-bearing  ores,  $1.75  is  a  moderate 
cost,  and  from  this  the  cost,  when  all  items  are  included,  may  rise 
to  $2.25  per  ton.  The  Allen-O 'Harra  automatic  furnace,  having  two 
straight  hearths  each  94  by  9  ft.,  and  resembling  the  Wethey  fur- 
nace, roasts  45  to  50  tons  daily  at  a  cost  of  78c.  per  ton.  The  Wethey 
furnace,  of  the  type  having  four  hearths,  each  65  by  10  ft.,  the  roast- 
ing proceeding  on  all  the  hearths,  roasts  90  tons  daily  to  5  to  6% 
sulphur,  at  a  cost  of  98c.  per  ton.  The  16-ft.  McDougall  furnace 
(Herreshoff  type),  having  five  hearths,  14%  ft.  diam.,  and  a  total 
area  of  830  sq.  ft.,  roasts  33  to  35  tons  daily  to  7%  sulphur  at  a  cost 
of  50c.  per  ton.  The  Bruckner  roasting  cylinder,  8Vi>  ft.  diam.  by 
22  ft.  long,  takes  a  charge  of  20  tons  (10  tons  daily),  and  in  48  hours 
roasts  it  to  4%  sulphur  at  a  cost  of  80c.  per  ton. 

It  will  be  noticed  that  the  low  cost  of  roasting  in  some  of  these 
furnaces  is  due  to  their  needing  no  fuel  after  coming  into  full  op- 
eration. To  obtain  this  effect  such  furnaces  have  several  hearths, 
and  are  compact.  On  account  of  this  compactness  they  lose  but 
little  heat  by  radiation. 

24.  BLAST  OR  POT-ROASTING  OF  ORES. 

Both  lead  and  copper  ores  are  treated  by  blast  or  pot-roasting, 
though  the  method  was  at  first  intended  for  lead-bearing  ores,  es- 
pecially for  galena.  We  have  alread.y  mentioned  the  difficulty  of 
roasting  galena  by  the  old  method,  in  the  reverberatory  furnace ;  but 
by  pot-roasting,  it  can  be  so  treated  as  to  remove  most  of  its  sulphur, 
with  less  loss  by  volatilization. 

Treatment  of  galena. — By  the  Huntington-Heberlein  process, 
called  also  the  '  II  and  H  process',  the  galena-bearing  ore  is  given  an 
incomplete,  rather  rapid  roast,  to  reduce  the  amount  of  sulphur  to 
12  to  14%.  The  product  from  the  roaster  is  mixed  with  a  certain 


OF    THE    COMMON    METALS.  113 

proportion  of  limestone  and  silicious  ore,  wet  down,  and  charged  into 
a  hemispherical  cast-iron  pot  81/?  ft.  diam.  by  4  ft.  deep,  having  a 
capacity  of  8  to  10  tons.  Within  the  pot,  and  forming-  a  false-bottom, 
is  placed  a  circular  arched  plate  perforated  with  %-in.  holes  to  admit 
air  to  the  charge  under  pressure.  Upon  the  false-bottom  is  scattered 
a  wheelbarrow-load  of  ashes,  then  a  carload  (one  ton)  of  hot  ore 
from  the  roaster.  On  this  is  dumped  8  tons  of  wet  charge.  Air, 
under  the  pressure  of  a  few  ounces,  is  admitted  beneath  the  false- 
bottom,  and  coming  up  through  the  hot  ore,  it  produces  a  burning- 
temperature  and  starts  the  combustion  of  the  charge.  The  heat 
gradually  ascending  to  the  top,  the  charge  becomes  red-hot,  and  S02 
and  SO3  escape.  At  the  end  of  the  roasting,  which  lasts  sometimes 
16  hours,  desulphurization  is  complete  and  there  remains  only  3  to 
5%  sulphur  if  the  charge  is  properly  burned.  The  pot  is  now  in- 
verted to  discharge  the  contents,  and  this  falls  out  in  an  agglome- 
rated, red-hot  mass.  It  is  broken  to  a  size  suited  to  subsequent 
treatment  in  the  blast-furnace. 

There  are  three  patented  variations  of  this  treatment  of  galena 
ores.  In  the  first,  the  Huntington-Heberlein  process  (above  de- 
scribed}, the  ore  is  mixed  with  limestone,  partly  roasted,  wet  down, 
charged  into  the  pots,  and  blown,  to  the  point  of  agglomeration.  In 
the  second,  the  Savelsburg  process,  the  ore  is  mixed  with  limestone 
and  charged  directly  into  the  pot  or  'converter',  the  preliminary 
roasting  being  omitted.  In  the  third,  the  Carmichael-Bradford  pro- 
cess, the  ore  is  mixed  with  gypsum  and  charged  into  the  pot  without 
preliminary  roasting. 

Treatment  of  copper  sulphides. — This  treatment,  like  the  Huiit- 
ington-Heberlein,  consists  in  blowing  the  partly-roasted  ore  or  matte 
in  pots,  but  in  this  case  no  lime  is  used.  The  charge  is  first  moist- 
ened so  that  it  easily  coheres.  The  minimum  quantity  of  water  to 
procure  the  best  result  is  3  to  4%  for  the  low-grade  matte,  4  to  6% 
for  high-grade  matte,  and  6  to  9%  for  ore.  In  the  case  of  ore  it  is 
found  that,  unless  the  water  is  in  excess,  ferric  oxide  is  produced, 
and  this,  forming  round  the  particles,  prevents  proper  slagging; 
whereas  with  sufficient  water,  ferrous  silicate  is  produced.  The 
charge  that  works  best  consists  of  about  one-third  of  pieces  1  to  l1^ 
in.  diam.,  and  two-thirds  of  fine  concentrate.  It  should  contain  15  to 
35%  Si02  and  15  to  25%  sulphur. 

In  treating  a  charge  the  false-bottom  is  covered  with  lumps  of 
roasted  ore,  after  this,  a  small  fire  of  chips  and  sawdust  is  started, 
and  urged  by  a  light  blast.  Ore  is  charged  next,  keeping  it  deeper 
at  the  sides,  until  the  pot  is  half  filled.  Holes  are  punched  into  the 


114  THE    METALLURGY 

mass  with  a  half-inch  pointed  rod,  and  through  these  appear  sul- 
phur vapor  and  sulphur  dioxide,  forced  by  the  blast.  In  about  an 
hour  a  ring  of  fire  begins  to  show,  the  remainder  of  the  charge  is 
then  put  on,  the  pot  is  covered  with  a  hood,  and  the  blast  gradually 
increased  to  13  oz.  per  sq.  in.  After  some  hours  the  evolution  of 
SO2  slackens  and  the  charge  gradually  becomes  red-hot  throughout. 
The  blast  is  then  stopped,  the  blast-pipe  disconnected,  and  the  con- 
verter inverted  to  discharge  the  contents.  As  arranged  in  some 
cases,  the  pot  with  its  load  is  picked  up  by  an  overhead  traveling- 
crane  and  dumped  near  a  large  crusher,  and  the  roasted  material 
crushed  into  lumps. 

The  time  required  to  treat  a  charge  varies  from  8  to  12  hours. 
The  product  consists  of  a  porous  sintered  mass  of  ferrous  silicate 
containing  shots  of  matte  and  free  silica.  It  is  well  suited  to  blast- 
furnace smelting.  .  A  large  quantity  of  the  product  in  one  instance 
contained  5.65%  sulphur,  while  as  little  as  5%  has  been  obtained  in 
the  treatment  of  ore.  The  process  is  particularly  suited  to  the  treat- 
ment of  matte,  to  which  15  to  25%  of  silicious  ore  is  added,  to  unite 
with  the  iron,  forming  ferrous  silicate.  In  ores  containing  pyrite, 
the  loosely  held  equivalent  of  sulphur  is  first  driven  off.  The  reac- 
tions should  proceed  rapidly  so  that  a  high  temperature  can  be  main- 
tained to  permit  the  formation  of  ferrous  silicate,  otherwise  ferric 
oxide  will  be  formed,  diminishing  the  amount  of  active  oxygen  near 
it.  Wherever  this  occurs,  the  temperature  drops,  and  a  patch  of 
unsintered  material  is  formed. 


PART  III.     GOLD 


PART  III.  GOLD. 

25.  GOLD  ORES. 

Gold  occurs  in  nature,  both  in  the  native  state  and  combined  with 
tellurium. 

Native  gold  occurs  in  vein-matter  disseminated  in  grains  or  par- 
ticles of  various  size,  and  it  is  found  not  only  in  quartz  veins  but  in 
veins  or  lodes  containing  hematite,  iron-pyrite,  arsenical-pyrite, 
blende,  and  galena.  In  pyrite  it  occurs  not  only  in  the  substance  of 
the  crystals  but  as  films  on  the  surface  of  these  crystals.  It  is  fre- 
quently accompanied  by  silver.  When  gold-bearing  veins  have  be- 
come disintegrated  and  swept  away  into  alluvial  deposits,  the  par- 
ticles of  gold,  where  released,  are  found  in  the  sand  and  gravel  of 
the  beds,  the  pebbles  and  boulders  themselves  (which  have  come  from 
the  country  rock),  being  in  general  barren  of  gold.  Gold  occuring 
in  this  way  is  called  alluvial  gold,  and  is  recovered  by  methods  of 
hydraulic  mining  or  dredging,  which  belong  to  mining  engineering 
rather  than  to  metallurgy.  We  shall  here  consider,  therefore,  the 
treatment  of  gold  ore. 

Gold  tellurides. — In  South  Dakota,  at  Cripple  Creek,  Colorado,  in 
Western  Australia  and  elsewhere  is  to  be  found  gold  combined  with 
tellurium  as  calaverite,  AuTe2  (containing  41.4%  Au  and  57.3%  Te)  ; 
also  gold  and  silver  combined  with  tellurium  as  sylvanite  (AuAg) 
Te2,  and  as  petzite  (Ag2Te,  Au2Te). 

Classification  on  treatment  basis. — Gold,  regarded  from  the  stand- 
point of  milling  and  amalgamation,  occurs  in  the  following  condi- 
tions : 

(1)  In  ordinary  amalgamable  form,  generally  called  'free-mill- 
ing'. 

(2)  In  some  form  of  intimate  physical  admixture,  with  other 
minerals,  or  in  chemical  combination  with  other  elements. 

(3)  As  rusty   gold,   sometimes  metallic   in  appearance   and  of 
usual  golden  color,  but  often  brown  and  lusterless.    In  this  supposed 
allotropic  form,  it  resists  the  action  of  mercury  in  the  process  of 
amalgamation. 


118  THE    METALLURGY 

26.     STAMP-MILL   AMALGAMATION. 

This  consists  in  crushing  the  gold-bearing  ore,  generally  in  a 
stamp-mill,  to  a  size  of  20-mesh,  or  finer  (using  6  to  8  tons  of  water 
to  one  ton  of  ore),  and  running  the  ore-pulp  over  plates  4%  ft.  wide 
by  6  ft.  long  to  which  the  gold  adheres  by  amalgamating  with  the 
mercury  with  which  the  plate  is  coated.  Occasionally  the  battery  is 
stopped  for  a  short  time,  and  the  gold-bearing  amalgam  is  scraped 
from  the  plate  and  treated  to  obtain  the  gold.  This  process  is  called 
'outside  amalgamation'.  Sometimes  amalgamated  plates  are  placed 
inside  the  mortar,  and  the  particles  of  gold  contained  in  the  ore  are 
driven  against  the  plates  by  the  movement  of  the  pulp  and  adhere. 
The  gold  is  recovered  in  the  same  way  as  on  the  outside  plates. 
Some  of  the  gold  particles  and  mercury  fall  into  the  crevices  be- 
tween the  dies,  at  the  bottom  of  the  mortar.  This  gold  is  recovered 
at  the  time  of  the  monthly  clean-up.  From  time  to  time  a  little 
mercury  is  placed  in  the  mortar  (about  1.5  oz.  per  ounce  of  gold  in 
the  ore),  as  the  crushing  proceeds. 

The  tailing,  or  residue  after  obtaining  the  gold,  may  be  run  to 
waste.  If,  however,  ore  contains  pyrite,  or  other  heavy  sulphide 
minerals  within  which  a  part  of  the  gold  is  locked,  only  a  part  of  the 
gold  can  be  recovered  by  amalgamation.  Pyrite,  being  heavy,  can 
be  caught  on  concentrating  tables,  and  the  concentrate  so  recovered 
can  be  sent  away  to  be  smelted  or  be  treated  by  leaching  methods. 
At  Treadwell  Island,  Alaska,  only  half  the  gold  in  the  ore  is  obtained 
by  amalgamation,  and  the  concentrate,  2^/2%  of  the  weight  of  the 
ore,  carries  most  of  the  remainder.  The  tailing  from  the  concentrat- 
ing tables  commonly  contains  a  small  amount  of  gold. 

The  mechanical  operation  of  pulverizing  the  ore  to  a  fineness  such 
as  to  liberate  the  gold  particles  from  the  enclosing  gangue  consists 
first  in  coarse-crushing  the  ore  in  a  Blake  type  crusher  (See  Fig. 
22),  which  discharges  into  an  inclined-bottom  bin,  and  withdrawing 
it  thence  to  be  crushed  in  a  stamp-battery. 

The  battery. — Fig.  44  is  a  perspective  view,  Fig.  45  a  side  eleva- 
tion, and  Fig.  46  a  front  elevation  of  a  10-stamp  battery  such  as  is 
used  in  gold  milling.  Fig.  45  is  a  good  example  of  a  dimensional 
drawing. 

Referring  to  Fig.  44,  the  parts  will  be  found  designated  as  fol- 
lows :  A,  mortar-block;  B,  mud-sills;  C,  cross-sills;  I),  side-posts;  E, 
platform ;  F,  G,  buck-staves ;  H,  lower  guide-timbers ;  7,  upper  guide- 
timbers  ;  J,  mortars ;  K,  screen ;  L,  die ;  M,  shoe ;  N,  boss  or  head ;  0, 
stem;  P,  tappet;  R,  cam-shaft;  S,  collars;  T,  cam-shaft  boxes;  U, 


OF    THE    COMMON    METALS. 


119 


cams ;  V,  cam-shaft  pulley ;  W,  line-shaft ;  X,  tightening-pulley ;  Y, 
water-pipes ;  Z,  automatic  feeder.  On  two  of  the  cross-sills  C,  is  to 
be  seen  the  frame  that  supports  the  apron-plate. 

The  mortar-blocks,  A,A,  are  made  of  timbers  set  on  end,  as  shown 
also  in  Fig.  45  and  46.     They  are  set  in  a  pit  that  extends  down  to 


Fig.   44.      PERSPECTIVE  VIEW  OF  TEN-STAMP  BATTERY. 

solid  rock  or  to  a  concrete  foundation,  and  upon  the  solidity  de- 
pends the  durability  of  the  battery.  They  are  bolted  together  hori- 
zontally, and  are  held  in  position  by  a  tamping  of  sand,  rock,  or  con- 
crete, completely  filling  the  pit  around  them.  Instead  of  wooden 
mortar-blocks,  concrete  ones  have,  of  late,  come  successfully  into 
use.  These  are  more  solid,  more  durable,  and  cheaper. 

The  mortar  J  is  firmly  held  to  the  mortar-blocks  by  long  bolts 


120 


THE    METALLURGY 


(See  Fig.  46),  three  thicknesses  of  blanket  of  a  piece  of  rubber-sheet- 
ing being  interposed  to  give  an  even  bearing. 


Fig.   45.      SIDE  ELEVATION  OF  TEN-STAMP  BATTERY. 


OF    THE    COMMON    METALS. 


121 


The  stamp-frames  consisting  of  the  mud-sills  B,  the  cross-sills  (7, 
the  side-posts  D,  the  buck-staves  F  and  G,  and  the  guide-timbers  H 
and  I  are  of  the  dimensions  shown  in  Fig.  45  and  46.  The  mud-sills, 
supporting  the  cross-sills  (7,  run  the  length  of  the  mill,  under  all  the 
batteries,  and  carry  the  line-shaft  W.  A  tightening  pulley  just  above 


-4- 


25'- 


Fig.  46.   FRONT  ELEVATION  OF  TEN-STAMP  BATTERY. 

the  lower  pulley,  serves  to  tighten  the  belt  and  set  the  battery  in  mo- 
tion. To  the  guide-timbers  H  and  /  are  bolted  the  guides,  often 
made  of  two  planks  each  4  by  14  in.  bored  vertically  with  3-in.  holes 
for  the  wrought-iron  or  steel  stems.  Otherwise  cast-iron  individual 
guides  are  employed. 

A  mortar,  as  seen  in  Fig.  47  and  48,  is  a  cast-iron  box  having  at 


122 


THE    METALLURGY 


one  side  a  feed-opening  through  which  the  ore  enters,  screen-open- 
ings at  one  or  both  sides  for  the  screens,  through  which  the  stamped 
ore  or  pulp  discharges,  a  heavy  base  on  which  rests  the  die,  and  sides 
enclosing  the  whole,  the  mortar  being  a  single  heavy  casting  weigh- 
ing 2  to  3  tons.  Fig.  47  is  a  mortar  of  the  single-discharge  type,  used 
in  gold-milling;  and  Fig.  48  a  double-discharge  mortar,  used  in  sil- 


\  I 

Fig.   47.      SINGLE-DISCHARGE   MORTAR. 

ver  milling.  With  the  single-discharge  mortar  the  output  is  less, 
and  the  pulp  is  retained  a  longer  time  in  the  mortar.  In  consequence 
it  is  more  finely  ground,  and  more  intimately  brought  in  contact  with 
the  inside  plates  when  used.  The  double-discharge  mortar  on  the 
other  hand  on  account  of  the  greater  total  opening,  discharges  freely, 
and  pulverizes  more  ore.  It  is  not  a  mortar  intended  for  amalgama- 
tion, but  for  crushing  only.  Upon  either  mortar  will  be  noticed  a 


OF    THE    COMMON    METALS. 


123 


deflecting-lip  at  the  bottom  of  the  feed-opening.  This  is  provided  to 
discharge  the  ore  nearer  the  center  of  the  mortar.  Beneath  it,  and 
protected  from  the  wear  of  the  entering  ore,  an  inside  plate  is  some- 
times placed. 

Screen-openings  are  provided  at  the  front,  and  in  the  double-dis- 
charge mortar  also  at  the  back  of  the  mortar.    The  screen  K,  Fig.  44. 


Fig.   48.      DOUBLE-DISCHARGE  MORTAR. 


tacked  to  a  wooden  frame,  is  made  either  of  wire-cloth  or  punched- 
plate.  Of  the  two  kinds,  the  punched-plate  screen  has  the  advant- 
age in  strength  and  first  cost.  The  wire-cloth,  on  the  other  hand,  if 
of  copper  or  brass,  is  durable  and  gives  a  greater  discharge  area. 
Thus  in  the  case  of  No.  7  punched-plate,  Fig.  49,  we  have  effective 
in  discharge  openings  but  10%  of  the  total  area ;  while  in  the  case  of 
the  wire-screen  27%  is  open  and  in  consequence,  there  is  less  sliming 


124 


THE    METALLURGY 


or  excessive  powdering  of  the  ore  because  of  its  prompt  escape  from 
the  mortar. 

To  increase  the  height  of  discharge,  when  the  pulp  is  to  be  re- 
tained a  longer  time  in  the  battery  for  finer  crushing  a  wooden 
chuck-block  is  placed  in  the  discharge-opening  beneath  the  screen- 
frame.  The  height  of  discharge  may  be  defined  as  the  vertical  dis- 
tance from  the  top  of  the  die  to  the  bottom  of  the  opening  of  the 
screen-frame.  Where  inside  amalgamation  is  practised,  an  inside 
amalgamated  plate  covering  the  chuck-block  is  used  in  addition  to 
other  plates,  and  on  these  plates  as  much  of  the  gold  as  possible  is 
collected.  Referring  to  Fig.  47  and  48,  we  see  a  foot-plate  that  cov- 
ers the  bottom  of  the  mortar,  upon  which  rest  the  five  cylindrical 
dies  8  to  9  in.  diam.  by  7  in.  high  when  new.  Resting  on  the  die  we 


Fig.    49.      PUNCHED   SCREEN. 


Fig.   50.      CANDA  CAM. 


notice  the  stamp-shoe  fitting  into  the  head,  which  in  turn  is  secured 
to  the  stamp-stem.  The  shank  of  the  shoe  is  made  a  little  smaller 
than  the  corresponding  recess  in  the  head,  and  in  inserting  it,  wooden 
shims  are  placed  around  the  shank,  and  the  stamp  is  dropped  upon 
it,  thus  wedging  the  shoe  into  the  head.  Dies  and  shoes  are  made 
of  chilled  cast-iron,  manganese  steel,  or  chrome-steel,  chilled  cast- 
iron  being  common.  The  wear  upon  shoes  and  dies  amounts  to  14 
to  114  lb.  per  ton  of  ore  crushed,  depending  upon  the  toughness  of 
the  ore.  On  an  average  these  parts  last  three  months.  In  the  case 
of  chilled  cast-iron,  %  lb.  at  6c.  per  lb.  or  4.5c.  per  ton  of  ore  treated 
is  an  average  cost  for  wear. 

The  tappets  P,  Fig.  44,  of  cast-iron  are  keyed  to  the  stems.  The 
stamps  are  lifted  8  to  16  in.  by  the  cams  H,  and  are  adjustable  to 
give  them  the  drop  desired.  At  the  same  time,  they  cause  partial 


OF    THE    COMMON    METALS.  125 

rotation  of  the  stem  on  its  axis  and  insure  an  even  wear  of  the  shoe 
and  die.  As  the  toe  of  the  cam  passes  on,  the  whole  'stamp'  (shoe, 
head,  stem,  and  tappet),  drops  with  the  impact  of  1000  lb.,  falling 
freely  upon  the  ore  in  the  mortar.  The  replacing  of  a  cam  of  the 
ordinary  type,  requiring  the  removal  of  the  key  that  holds  it,  is  a 
tedious  operation,  and  to  overcome  this,  a  self-tightening  cam  like 
the  Canda,  Fig.  50,  has  been  devised.  This  consists  of  a  curved  tap- 
ering key  that  fits  an  eccentric  recess  in  the  cam.  Into  the  shaft  is 
set  a  slightly  projecting  pin  which  engages  the  key  in  a  groove  or 
recess  so  that  when  cam  and  key  are  set  upon  the  shaft  over  the  pin 
and  turned,  the  key  wedges  and  ^tightens  the  cam  to  the  shaft.  The 
cams  are  set  on  the  cam-shaft  at  various  angles  so  that  the  stamps 
drop  at  regular  intervals  and  in  a  pre-determined  order.  The  stamps 
being  numbered  consecutively  from  left  to  right,  in  a  5-stamp  bat- 
tery, a  favorite  order  of  drop  is  No.  1,  47  2,  5,  3. 

The  cam-shaft  is  carried  by  boxes  secured  to  the  side-posts,  and  is 
driven  by  a  large  pulley  made  of  wood  in  preference  to  iron,  in  order 
to  withstand  the  shock  of  operation.  When  it  is  desired  to  stop  the 
battery,  the  stamps  are  'hung-up'  by  inserting  under  the  tappets  the 
'fingers',  one  of  which  is  seen  in  the  elevation,  Fig.  45. 

Ore  is  fed  to  the  battery  automatically  by  a  feeder  marked  Z,  in 
Fig.  44  and  45,  that  takes  its  supply  from  the  sloping-bottom  stor- 
age-bins above.  As  seen  in  Fig.  47  and  48,  there  is  a  ledge  or  lip  on 
the  mortar  over  which  the  pulp  flows  to  the  apron-plate.  In  the 
double-discharge  mortar,  the  pulp-flow  at  the  back  of  the  mortar 
joins  that  at  the  front  by  passing  through  a  hole  or  passage  in  the 
base  of  the  mortar. 

27.     OPERATION  OF  THE  STAMP-BATTERY. 

From  the  storage-bin,  indicated  behind  the  battery  in  Fig.  44, 
and  shown  in  section  in  Fig.  45,  the  ore,  crushed  to  l^-in.  size,  passes 
a  chute  to  the  hopper  of  the  automatic-feeder.  A  feeder  of  the  Chal- 
lenge type  is  shown  in  Fig.  51. 

At  each  drop  of  one  of  the  stamps,  a  collar  on  the  stem  strikes 
the  end  of  the  projecting  horizontal  lever,  and  this  lever  actuates  the 
feed-plate,  revolving  it  slowly  against  a  fixed  scraper,  and  causing 
ore  to  fall  into  the  feed-opening  of  the  mortar.  As  the  ore  accumu- 
lates under  the  shoes,  the  stroke  shortens,  and  the  thrust  of  the  hori- 
zontal lever  being  less  the  feed  corresponding  is  lessened.  As  the  ore 
supply  lessens  the  stroke  lengthens.  Thus  the  action,  according  to 
the  needs  of  the  battery,  is  automatic.  Water  is  introduced  through 
a  pipe  at  the  feed-opening,  and  mixing  with  the  pulverized  ore, 


126 


THE    METALLURGY 


splashes  out  through  the  screen  at  each  fall  of  the  stamps.  The  feed 
is  regulated  so  as  to  cause  the  stamp  to  strike  with  a  sharp,  hard 
blow,  but  with  little  of  the  rebound  that  would  occur  with  too  thin 
a  layer  of  ore. 

Mercury  fed  to  the  battery.— This  varies  from  1  to  6  oz.  per  ounce 
of  gold  caught,  the  average  being  1.5  oz.  Added  a  little  at  a  time  in- 
side the  mortar,  it  works  out  in  part  upon  the  apron-plates.  As  to  the 


Fig.    51.      AUTOMATIC   FEEDER. 


amount  to  use,  a  safe  guide  is  the  appearance  of  the  plates.  If  they 
are  hard  the  indication  is  insufficient  mercury ;  if  mercury  is  dis- 
tinctly visible  on  them,  either  in  patches  or  in  streaks,  too  much  is 
being  added.  The  mercury  should  be  free  from  base  metals  that 
cause  it  to  'sicken',  or  break  into  coated  globules.  Such  globules 
refuse  either  to  coalesce  or  adhere  to  the  amalgamated  surface,  and 
are  swept  away  with  the  pulp.  Mercury  is  best  that  already  con- 
tains gold  and  silver. 


OF    THE    COMMON    METALS.  127 

Dressing  the  plates. — The  outside  or  apron-plates  are  dressed 
three  or  four  times  daily,  the  operation  taking  about  15  minutes.  To 
do  this,  the  feeding  is  stopped  to  permit  the  ore  to  work  out  of  the 
mortar,  the  stamps  are  'hung-up',  and  the  apron-plate  washed  clean 
with  a  stream  of  water.  A  rubber-edged  scraper,  resembling,  but 
heavier  than  a  window-cleaner,  is  used  to  scrape  the  plate.  The 
amalgam,  perhaps  half  a  pint  in  quantity,  is  scraped  together  with 
this,  gathered  up,  and  placed  in  an  enameled  cup.  If  the  surface  is 
too  hard  for  the  scraper,  the  amalgam  is  softened  by  sprinkling  a 
little  mercury  upon  it.  Dressing  being  completed,  the  stamps  are 
started,  and  feeding  is  resumed. 

Plates  are  liable  to  become  tarnished  with  salts  of  copper  that 
form  a  coating  like  verdigris  upon  the  surface.  Since  the  tarnished 


Fig.    52.      MERCURY   TRAP. 

spot  collects  no  gold,  the  stains  must  be  removed.  To  do  this  a  solu- 
tion of  sal-ammoniac  is  applied  to  the  stained  parts  with  a  scrubbing 
brush  when  the  battery  is  stopped.  In  a  few  minutes  this  chemical 
is  washed  off,  and  potassium  cyanide  and  then  mercury  are  rubbed 
on,  and  the  plate  immediately  washed  clean.  Apron-plates  have  a 
grade  of  0.5  to  1.75  in.  per  foot,  ore  containing  sulphides  requiring 
the  steepest  grade.  When  the  pulp  is  flowing  over  the  plate  in  a 
proper  manner  it  travels  down  in  the  form  of  series  of  ripples  or 
waves,  bringing  the  particles  of  gold  in  contact  with  it. 

To  save  escaping  particles  of  amalgam  or  globules  of  mercury 
that  fail  to  adhere  to  the  plate,  a  mercury  trap,  Fig.  52,  is  provided. 
This  is  especially  important  where  no  concentration  of  the  tailing  is 
attempted.  In  shape  it  is  an  inverted  frustum  of  a  pyramid.  The 
pulp  enters  by  the  vertical  pipe,  and  escapes  over  the  wooden  block 
shown  belted  to  the  side  of  the  trap.  The  pulp  overflows  through 
the  pipe  at  the  side.  Some  heavy  sulphides  accumulate  in  this,  but 


128  THE    METALLURGY 

the  amalgam  works  down  into  the  bottom.  Occasionally  the  accumu- 
lation at  the  bottom  of  the  trap  is  emptied  through  the  plug-hole  in 
the  bottom,  and  is  panned  to  recover  the  amalgam. 

The  clean-up.— In  operating  a  40-stamp  mill,  it  is  customary  to 
remove  the  entire  accumulation  of  amalgam  from  the  mortar  every 


Fig.    53.      CLEAN-UP   PAN. 

two  to  four  weeks,  and  to  thoroughly  dress  and  scrape  the  plates. 
To  do  this,  two  batteries  (10  stamps),  at  a  time  are  hung  up.  The 
screens,  inside  plates,  and  dies,  are  taken  out,  and  the  contents  of  the 
mortar,  perhaps  two  or  three  bucketfuls,  is  carefully  scraped  out  and 
fed  to  the  next  batteries.  The  plates  are  then  dressed,  the  dies  and 


OF    THE    COMMON    METALS.  129 

screens  returned  to  place,  and  the  two  batteries  again  started.  The 
next  two  are  taken  in  the  same  way,  and  so  on.  Finally  the  last 
batteries  are  hung  up,  the  contents  removed  and  put  into  a  clean-up 
pan,  Fig.  53,  with  the  amalgam  from  the  well-scraped  plates.  Three 
men  can  'clean  up'  a  40-stamp  mill  in  five  to  seven  hours. 

The  clean-up  pan,  Fig.  53,  3  ft.  diam.  by  3  ft.  deep,  making  12  to 
15  r.p.m.,  is  used  for  grinding  the  sand,  pyrite,  fragments  of  iron, 
and  other  substances,  collected  with  the  amalgam  in  the  battery 
clean-up.  The  charge,  of  300  lb.,  mixed  with  water,  partly  fills  the 
pan.  It  is  ground  to  a  fine  mud  in  3  to  4  hours,  during  which  time 
50  Ib.  mercury  is  added,  and  the  mixing  continued  a  few  hours 
longer.  The  pulp  is  then  diluted  with  water,  the  muddy  portion  de- 
canted by  taking  out  a  plug  that  is  a  little  above  the  bottom.  The 
residual  mercury  and  amalgam,  with  some  of  the  mud,  is  withdrawn 
through  the  lowest  plug-hole,  panned,  treated  in  an  enameled-ware 
bowl  with  nitric  acid,  and  well  washed  until  clean.  The  residual 
mercury  and  amalgam  is  strained  through  chamois  skin,  or  through 
canvas,  to  remove  the  excess  of  mercury.  Gold  amalgam,  when  well 
squeezed  through  cloth,  contains  35  to  45%  gold.  The  mercury  that 
has  been  removed  by  filtration  still  retains  0.5  per  cent. 

Retorting. — In  the  smaller  gold-mills,  amalgam  is  retorted  in  a 
pot-shaped  retort,  Fig.  54.  In  larger  mills  a  horizontal  cylindrical 
retort  is  used,  though  the  latter  is  chiefly  for  silver-mills  where  a 
large  quantity  of  amalgam  is  made.  The  retort,  Fig.  54,  is  filled  two- 
thirds  full  of  amalgam,  placed  in  a  wind-furnace  and  supported  by 
perforated  cast-iron  thimble  that  rests  on  the  grate-bars.  The  cover 
is  luted  and  tightly  clamped.  A  pipe  through  which  the  mercury 
vapor  passes  leads  out  from  the  cover  and  turns  downward.  The 
outer  leg  of  the  pipe  is  water-cooled  and  the  end  dips  into  a  tub  of 
water.  A  fire  having  been  started  on  the  grate-bars,  the  retort 
gradually  heats  until  mercury  vapor  begins  to  come  over.  It  con- 
denses in  drops  in  the  cool  pipe  and  collects  in  the  tub  below.  When 
a  distilling  heat  is  obtained,  the  fire  is  checked  and  the  retort  kept 
at  an  even  temperature  for  one  or  two  hours,  after  which  it  is  heated 
to  redness  to  expel  the  last  of  the  mercury. 

The  mercury,  collected  by  condensation  in  the  tub  under  water, 
is  used  again.  This  accounts  for  the  small  loss,  which  averages  in 
California  practice  but  0.5  oz.  per  ton  of  ore  treated.  Mercury  is 
lost  by  'flouring'  and  by  'sickening'.  The  first  of  these  losses  is  in- 
dicated by  a  white  appearance,  and  is  caused  by  excessive  agitation 
in  the  air,  which  breaks  it  into  globules  or  particles  so  fine  as  not 
again  to  unite  or  at  least,  not  without  great  difficulty.  Sickened 


130 


THE    METALLURGY 


mercury  is  black,  and  owes  its  appearance  to  the  presence  of  base 
metals,  as  already  explained.  The  retorted  residue,  still  containing 
0.5  to  \%  mercury,  is  porous,  and  consists  of  gold  from  500  to  900 
fine.  It  is  melted  in  a  plumbago  crucible  in  a  wind-furnace  with 
soda  and  borax,  and  when  it  contains  base  metal,  with  a  little  nitre 


which  serves  to  toughen  it.    The  melt  is  poured  into  an  ingot  mold ; 
and  the  bar,  cleaned  from  adhering  slag,  is  shipped  to  the  mint. 

28.  GENERAL  ARRANGEMENT  OF  A  GOLD  MILL. 

Fig.  55  is  a  plan,  and  Fig.  56  an  elevation  of  a  20-stamp  mill, 
where  ore  is  amalgamated  and  the  tailing  concentrated. 


OF    THE    COMMON    METALS. 


131 


The  ore  enters  the  mill  on  a  high-level  track  in  tram-cars  and  is 
dumped  into  a  gyratory  crusher  and  broken  to  1  to  l^-in.  size. 
From  the  crusher  it  goes  by  a  shoot  to  either  of  two  storage-bins  for 
the  two  batteries  of  10  stamps  each.  The  crushing  is  done  during 
the  10-hour  day-shift,  and  the  bins  are  large  enough  to  hold  a  day's 
supply.  From  these  bins  the  ore  discharges  into  automatic  feeders 
(not  shown) ,  and  from  there,  in  constant  supply  to  the  stamp  bat- 


Fig.   55.      PLAN  OF  TWENTY-STAMP  MILL. 

teries.  The  batteries  are  driven  by  a  direct-coupled  driving  shaft 
supported  upon  the  mud-sills  of  the  batteries,  as  shown  by  IF,  Fig. 
44.  The  pulp,  splashing  through  the  battery  screens,  flows  over  a 
set  of  short  apron-plates.  The  tailing  from  the  plates  unites  in  a 
launder,  and  finally  falls  into  a  distributing  box  that  commands  the 
tables.  A  distribution  is  made  here  and  one-fourth  is  supplied  by 
a  launder,  to  each  vanner.  The  tailing  from  the  vanners  is  wasted. 
The  concentrate  is  collected,  and  shipped  for  smelting.  The  method 


132 


THE    METALLURGY 


of  driving  the  machines  is  indicated  in  Fig.  56.  Above  the  battery 
runs  an  over-head  track,  carrying  a  trolley  and  heavy  chain  tackle, 
by  means  of  which  a  stamp  or  any  heavy  part  can  be  removed  readily 
or  replaced. 

When  ore  contains  heavy  sulphides,  such  as  pyrite,  the  tailing 
from  the  plates  is  concentrated  on  the  Frue  vanner,  or  the  Wilfley 
concentrating-table.  Thus  a  valuable  product  is  obtained  in  a  small 
bulk,  and  any  escaping  particles  of  amalgam  are  caught  with  the 


Fig.    56.      ELEVATION  OF  TWENTY-STAMP  MILL. 

concentrate.  When  ore  contains  gold,  both  coarsely  and  finely  dis- 
seminated, the  usual  method  has  been  to  recover  the  coarse  gold  by 
amalgamation,  and  treat  the  residue  by  cyanidation  to  extract  the 
gold  not  recovered  by  amalgamation.  Another  practice,  dispensing 
with  amalgamation,  has  been  to  crush  with  stamps,  separate  the 
slimed  material,  and  re-crush  the  coarser  portion  (the  sand),  thereby 
finely  grinding  the  auriferous  particles  so  that  they  may  be  cyanided 
in  a  reasonable  time,  not  possible  with  ore  consisting  of  coarse  auri- 
ferous particles. 


OF    THE    COMMON    METALS.  133 

29.     CALIFORNIA  AND  COLORADO  PRACTICE  IN  GOLD- 
MILLING. 

For  free-milling  ores,  where  the  gold  is  coarse,  and  where  in  con- 
sequence fine  grinding  is  not  needed  to  release  the  gold  from  its 
matrix,  California  practice,  with  rapid  drop  and  low  discharge  is 
preferred,  since  in  this  way  large  tonnage  is  secured.  On  the  other 
hand  in  Gilpin  county,  Colorado,  the  ore  contains  10%  gold-bearing 
pyrite  in  which  the  gold  is  finely  disseminated.  To  release  the  gold, 
the  ore  must  be  longer  retained  within  the  mortar  and  finely  crushed. 
While  retained  there  the  fine  float-gold,  set  free,  has  time  to  come  in 
contact  with  the  inside  amalgamating  plates.  These  recover  about 
75%  of  the  gold. 

In  California  practice,  the  stamps  weigh  900  to  1100  lb.,  drop  80 
to  110  times  per  minute,  and  have  a  fall  of  5  to  9  in.  In  Colorado 
practice  the  stamps,  weighing  600  to  800  lb.r  drop  only  25  to  30 
times  per  minute,  but  through  a  height  of  18  to  20  in.  In  general, 
the  greater  number  of  drops  per  minute  the  greater  is  the  tonnage. 
To  pulverize  the  ore  finer  and  retain  it  in  the  battery  a  longer  time 
for  inside  amalgamation,  a  high  discharge  is  needed,  and  this  is  ob- 
tained by  using  a  deeper  chuck-block,  bringing  the  height  in  the 
Colorado  type  of  battery  to  13  to  18  in.,  while  in  California  practice 
it  is  but  5  in.  The  objection  to  a  high  discharge  is  that  more  of  the 
ore  is  slimed  and  is  difficult  to  concentrate,  causing  a  high  loss  of 
gold-bearing  pyrite  in  the  tailing.  In  California  practice  4  tons  per 
head  may  be  crushed;  in  the  Colorado  practice  it  is  but  1  to  1.5  tons. 
We  thus  have  the  following  practices  for  average  ore : 


Drops  per  minute  

California. 
95 

Colorado. 

28 

Height  of  drop    (inches)    

7 

18 

Height  of  discharge   (inches) 
Weight  of  stamps  (pounds)  

5 

1000 

15 
700 

Actual  horse-power  per  stamp 
Capacitv  in  24  hours  (tons)   . 

2.02 
4 

1.07 
1.25 

The  theoretical  horse-power  of  a  single  stamp  is  calculated  by 
multiplying  the  weight  in  pounds  by  the  distance  lifted  per  minute, 
in  feet,  and  dividing  by  33,000.  The  actual  horse-power  may  be 
computed  as  1.2  times  the  theoretical.  The  harder  and  tougher  the 
ore,  the  slower  will  be  the  crushing;  while  the  softer,  and  more  fri- 
able the  ore  and  the  coarser  the  product,  the  larger  will  be  the  ton- 
nage. 


134  THE    METALLURGY 

30.     COST  OF  GOLD-MILLING. 

In  South  Africa,  where  amalgamation  has  been  followed  by  cy- 
anidation  of  the  tailing,  the  cost  of  milling  and  amalgamation  has 
been  $0.72  to  $1.20  per  ton  and  $0.96  to  $1.44  per  ton  additional  for 
cyaniding.  These  costs  are  based  upon  an  output  of  4%  to  5  tons 
per  stamp. 

In  California  in  1896  at  a  30-stamp  mill  there  were  crushed  and 
concentrated  33,512  tons  of  ore  at  the  following  itemized  cost  per 
ton: 

Shoes  and  dies   $0.029 

Screens    0.003 

Mercury    0.007 

Hardware,  belting,  and  firewood  0.021 

Water  for  power   0.095 

Freight,  cyanide,  oil,  and  grease 0.006 

Lumber    0.008 

Miscellaneous    0.007 

Assay  and  office  supplies    0.008 

Silver-plated   plates    0.007 

Water-pipes  and  connections   0.021 

Hauling  sulphides   0.020 

Express  on  bullion   0.006 

Taxes  and  insurance 0.010 

Superintendence  and  labor  0.160 


$0.408 

A  summary  shows  that  of  this  cost,  $0.153  was  for  repairs,  $0.16  for 
labor,  and  $0.095  for  power. 

At  a  certain  30-stamp  mill  in  Idaho,  the  cost  was  $0.41  per  ton ; 
at  a  40-stamp  mill  in  California,  $0.50,  and  at  another  $0.49  per  ton. 
At  a  40-stamp  gold-mill  in  Gilpin  county,  Colorado,  the  cost  was  $0.84 
and  at  another  $1.47  per  ton.  Olcott  gives  the  cost  of  milling  at  sev- 
eral California  mills  as  varying  from  $0.20  to  $0.75  per  ton,  and  in 
a  Gilpin  county,  Colorado,  mill  $0.95  per  ton. 

31.     THE  HYDROMETALLURGY  OF  GOLD. 

At  the  present  time  there  are  two  methods  by  which  gold  is  dis- 
solved from  its  ore  by  chemical  solvents.  In  either  process  the  first 
step  is  to  obtain  the  gold  in  aqueous  solution,  then  to  precipitate  it 
from  the  clear  filtrate,  and  finally  to  get  it  in  the  form  of  a  bar  or 
ingot. 


OF    THE    COMMON    METALS.  135 

These  two  processes  are:  (1)  The  chlorination  or  Plattner 
process,  by  which  the  gold  is  obtained  in  solution  as  a  chloride  by 
the  action  of  an  aqueous  solution  of  chlorine  gas.  (2)  The  cyanide 
or  Mac  Arthur-Forrest  process  in  which  the  solution  of  the  gold  is 
effected  by  a  weak  cyanide  solution,  the  dissolved  gold  then  being 
present  as  potassium  auro-cyanide.  With  certain  refractory  ores, 
the  activity  of  the  solution  is  greatly  increased  by  the  use  of  bromine 
or  bromo-cyanogen  in  addition  to  the  potassium  cyanide. 

Extraction  of  gold  by  means  of  a  solvent  in  aqueous  solution  is 
also  practised  where  gold  cannot  be  completely  extracted  by  amal- 
gamation. This  often  is  the  case  with  pyrite  ores;  and  extraction 
can  be  practised  to  advantage  not  only  where  amalgamation  is  un- 
suitable but  where  smelting  is  expensive. 

Gold  in  ore  occurs  in  particles  of  various  sizes,  both  as  grains 
readily  seen,  and  in  particles  of  microscopic  size.  When  the  particles 
are  visible,  or  when  the  ore  shows  'colors'  upon  panning,  the  gold  is 
called  coarse,  and  such  particles  generally  can  be  recovered  by 
amalgamation.  Gold  often  occurs  in  finely  disseminated,  microscopic 
particles,  not  visible  to  the  eye,  and  in  films  on  the  surface  of  pyrite 
crystals.  If  the  ore  can  be  ground  so  fine  as  to  unlock  the  crystals,  or 
if  it  is  permeable  to  solutions,  gold  can  be  dissolved  in  aqueous  sol- 
vents, such  as  chlorine  or  potassium  cyanide.  Advantage  is  taken 
of  the  solubility  of  the  released  gold  particles,  and  leaching  or  per- 
colation methods,  in  tanks  or  vats,  are  practised  with  this  in  view. 
The  solution  soaks  through  the  ore4  comes  in  contact  with  gold  par- 
ticles, and  dissolves  them,  or  by  another  process,  the  finely-ground 
ore  or  slime  is  agitated  with  the  solution,  and  the  pulp  is  filtered  and 
washed  in  filter-presses.  The  clear  filtrate,  in  any  case,  is  treated 
by  a  suitable  precipitant  to  obtain  the  gold  in  small  bulk,  and  the 
precipitated  gold  is  melted  and  cast  in  the  form  of  a  bar  or  ingot 
for  sale. 

There  are  thus  three  stages  in  any  method  of  extracting  gold  by 
aqueous  solvents:  (1)  The  ore  is  finely  ground,  and  when  refrac- 
tory roasted,  to  convert  the  gold  into  a  soluble  form,  and  render  it 
accessible  to  the  solution.  (2)  The  gold  is  extracted  from  the  ore 
by  means  of  a  dilute  solvent,  using  a  tank  with  a  filter-bottom,  or 
agitating  the  ore,  pulverized  to  a  thin  pulp,  using  a  filter-press  for 
the  separation  of  the  solution.  (3)  The  gold  in  the  solution  is  pre- 
cipitated (a),  in  chlorination  by  hydrogen  sulphide  or  other  precipi- 
tating agent  or  (b),  in  cyanidation  by  the  use  of  zinc-shaving  or 
dust.  The  precipitate  is  collected,  dried,  and  melted  into  an  ingot. 

The   cyanidation  has  proved   a  remarkably  cheap  and  efficient 


136  THE    METALLURGY 

method  of  extraction,  but  it  has  limitations,  not  only  in  respect  to 
the  solubility  of  the  gold,  but  because  of  the  interference  of  com- 
pounds that  sometimes  are  present,  notably  those  of  copper,  that 
interfere  with  extraction  in  various  ways.  The  process  has  the 
advantage  over  the  chlorination  method  in  that  silver,  as  well  as 
gold,  can  be  extracted.  Under  favorable  conditions  the  extraction 
is  high,  and  modern  methods  have  reduced  the  cost  of  treatment  to 
a  low  figure. 

Pyrite  ore,  exposed  to  the  weather,  becomes  acid  in  reaction,  and 
if  treated  by  cyanide,  decomposes  and  destroys  the  potassium  cyan- 
ide. To  correct  this,  ore  is  first  treated  by  a  wash  of  dilute  caustic 
soda,  or  if  acid,  mixed  with  caustic  lime  in  sufficient  quantity  to 
overcome  acidity. 

When  ore  is  refractory  and  requires  preliminary  roasting,  the 
cost  of  roasting  adds  much  to  the  cost  of  treatment.  In  chlorination, 
roasting  is  always  necessary,  and  in  any  case  it  improves  the  condi- 
tion of  the  ore  and  makes  it  porous  and  permeable  when  leached  or 
filter-pressed. 

32.     CHLORINATION  OF  GOLD  ORES. 

This  consists  in  attacking  the  gold  of  the  ore  with  chlorine  to 
form  the  soluble  gold  chloride,  and  dissolving*  out  the  gold  chloride 
in  water. 

The  complete  process  consists  of  the  following  parts:  (1)  Pre- 
paration.— Crushing  and  roasting  the  ore.  (2)  Extraction. — Bring- 
ing gold  into  the  form  of  gold  chloride,  which  is  soluble  in  water, 
and  leaching  this  out,  obtaining  a  clear  filtrate  that  contains  the 
gold  chloride.  .(3)  Segregation.— Precipitating  the  gold  from  the  fil- 
trate in  metallic  form,  or  as  a  sulphide  and  collecting  and  refining  the 
precipitate  to  obtain  the  gold  in  the  form  of  an  ingot. 

Ores  suited  to  chlorination. — An  ideal  ore  for  chlorination  is  one 
in  which  the  gold  is  present  in  a  state  of  division,  in  which  bases  are 
absent  that  would  be  attacked  by  chlorine,  and  silver  if  present  is 
in  such  a  condition  as  not  to  coat  the  particles  of  gold  with  insoluble 
silver  chloride.  While  the  cyanide  process  is  better  for  the  treat- 
ment of  low-grade  ores,  many  refractory  high-grade  ores  have  given 
better  results  by  chlorination. 

Ores,  in  which  the  gangue  consists  of  hydrated  iron-oxide,  are 
extremely  difficult  to  amalgamate.  Not  only  is  the  gold  finely  di- 
vided, but  the  ore  is  slimy  and  forms  a  coating  on  the  amalgamat- 
ing-plates.  Such  ores  give  satisfactory  results  by  barrel-chlorina- 
tion.  Silver  is  not  recovered  by  chlorination  since  it  becomes  an  in- 


OF   THE    COMMON    METALS.  137 

soluble  silver  chloride.  If,  however,  sufficient  silver  be  present  to 
pay  the  increased  cost,  salt  may  be  used  in  roasting  and  the  silver 
extracted  by  means  of  sodium  hyposulphite.  A  recovery  of  60%  of 
the  silver  is  possible  in  this  way. 

Below  are  analyses  of  ores  that  have  been  successfully  treated  by 
chlorination : 

Fig.    57.      TABLE  OF  GOLD  ORES. 

(1)  (2)  (3)  (4) 

Delano  mine,   Eureka  and 

Portland  mine,  Boulder  Idaho  mines, 

Cananea,  Cripple  Creek,  county,  Grass  Valley, 

Mexico.  Colo.  Colo.  Cal. 

Cu    0.10  0.85 

Zn    0.78  

Pb    0.78 

Mn  and  Fe 3.40  4.15  6.00  40.65 

S     0.80  2.49  2.20  32.80 

SiOa    83.50  54.91  89.50  12.64 

A1,0S    3.20  17.80               0.19 

MgO   and   CaO 2.40  0.25  5.83 

Alkalis 12.00              

Ag  (oz.  per  ton) 1.35  0.50  1.25  2.00 

All  (oz.  per  ton) 1.23  1.00  0.65  8.00 

Roasting. — Oxidized  ore  at  Mount  Morgan,  Western  Australia, 
containing  but  a  trace  of  sulphide,  is  crushed  in  Krupp  ball-mills 
and  quickly  roasted  (flash-roasted),  in  a  cylindrical  roaster  like  the 
White-Howell,  Fig.  40,  to  dehydrate  it  and  make  it  porous. 

Ore,  containing  sulphur,  arsenic,  and  antimony,  is  crushed  to  10 
to  30-mesh  size,  and  is  roasted  to  expel  these  elements,  to  oxidize  the 
bases,  to  leave  the  gold  in  such  form  as  to  be  attacked  by  chlorine, 
and  to  make  the  ore  porous,  accessible  to  chlorine,  and  more  easily 
leached. 

In  attempting  to  chlorinate  ore  (1)  of  the  table  unroasted,  the 
sulphur  (0.8%),  consumed  chlorine,  and  an  extraction  of  only  25% 
resulted.  After  dead-roasting,  98.4%  of  the  gold  was  extracted. 

Cripple  Creek  ore  (2)  containing  2  to  3.5%  sulphur  was  roasted 
to  0.08  to  0.10%,  then  cooled,  and  chlorinated.  The  extraction  of 
gold  was  92  to  95%  of  the  amount  present.  The  loss  of  gold  in  roast- 
ing, due  to  volatilization  and  dusting,  is  commonly  3  per  cent. 

Ore  (3)  containing  gold  telluride,  is  broken  by  graded  crushing 
(See  Fig.  25),  to  20-mesh  size,  and  is  roasted  in  a  Pearce-turret 
roaster.  (See  Fig.  40  and  41).  The  ore,  discharged  from  the  furn- 
ace, passes  to  an  automatic  cooling-device,  consisting  of  vertical 
tubes  surrounded  by  water.  The  ore  passes  down  the  tubes,  and  is 
gradually  removed  in  cooled  condition  at  the  bottom. 

Ore  (4)  is  a  concentrate  from  gold-milling,  containing  much  sul- 
phide, and  typical  of  California  ore  to  which  chlorination  is  applied. 


138  THE   METALLURGY 

The  coarse  gold  has  been  removed,  by  milling  and  amalgamation, 
and  the  concentrate,  generally  1.5  to  2%  the  weight  of  the  ore  milled, 
contains  gold  in  fine  particles.  The  concentrate  is  roasted  generally 
in  long-bedded  reverberatory  furnaces  (Fig.  32),  60  ft.  long,  and  3 
tons  capacity  per  24  hours.  This  ore  contains  copper,  lead,  lime,  and 
magnesia,  all  of  which  consume  chlorine,  and  form  chlorides.  To 
prevent  this,  it  has  been  customary  to  add  salt,  to  the  extent  of  0.75 
to  1.5%  of  the  charge,  at  or  near  the  completion  of  the  roast.  If 
roasting  has  been  thorough  up  to  this  time,  copper  is  present  as  CuO, 
lead  as  PbS04,  lime  as  CaO,  and  magnesia  as  MgO. 

Were  the  copper  present  at  the  end  as  CuS04  it  would  react  with 
the  salt  forming  a  chloride  of  copper.  The  common  salt  also  reacts 
upon  the  gold  and  forms  gold  chloride.  Both  these  chlorides  are 
volatile,  and  the  CuCl,  in  volatilizing,  promotes  the  volatilization  of 
the  gold.  Professor  S.  B.  Christy,  who  conducted  muffle  roasting- 
tests  on  pyrite  mixed  with  5%  of  salt,  found  at  a  dull-red  heat,  12% 
of  the  gold,  and  at  a  cherry-red  21%,  to  be  lost.  The  losses  in 
silver  were  somewhat  less,  being  7%  and  17%  under  these  conditions. 
As  long  as  sulphur  is  present  it  protects  the  gold  from  attack,  but 
when  sulphates  have  been  formed,  and  are  causing  the  abundant 
evolution  of  chlorine  by  reaction  with  the  salt,  the  escaping  gas 
carries  gold  chloride,  the  gold  being  unprotected  by  sulphur  at  the 
time  of  chloridization. 

Lead  sulphate  similarly  reacts  with  salt,  forming  lead  chloride, 
which  does  not  consume  chlorine.  When  much  lead  is  present,  how- 
ever, it  is  removed  by  leaching  with  hot  water  before  treating  with 
chlorine.  Lime  and  magnesia  are  converted  by  the  salt  into  chlorides, 
and  in  this  form  consume  no  chlorine.  The  process  of  roasting  is 
therefore  conducted  as  follows : 

The  ore  is  thoroughly  roasted  at  a  low-red  heat.  The  tempera- 
ture is  then  a  bright-red  (850°C.),  to  decompose  copper  sulphate. 
The  salt  is  then  added  and  thoroughly  incorporated,  and  the  tem- 
perature reduced  to  prevent  volatilization  of  the  gold.  The  quantity 
of  salt  to  be  added,  the  time  needed  for  roasting,  and  the  tempera- 
ture compatible  with  the  minimum  loss  of  the  gold,  should  be  de- 
termined experimentally  for  each  kind  of  ore. 

33.     EXTRACTION  OF  THE  GOLD  BY  CHLORINATION. 

There  are  two  methods  of  chlorinating  ore.  These  are  (1)  the 
vat,  or  Plattner  process.  (2)  The  barrel,  or  Theiss  process.  In  the 
vat  method  the  ore,  after  moistening,  is  charged  into  the  vat  and 
subjected  to  the  action  of  chlorine  gas  conducted  into  it  from  a  sepa- 


OF   THE    COMMON    METALS. 


139 


rate  vessel.  No  motive  power  is  required,  and,  aside  from  the  cost 
of  the  roasting-furnace,  but  small  investment  is  involved  in  the  plant. 
The  process  is  suited  to  the  treatment  of  a  small  daily  supply  of 
concentrate.  In  barrel  chlorination,  ore  is  charged  into  lead-lined 
steel  barrels  or  cylinders,  and  there  exposed  to  the  action  of  chlorine 
generated  from  chemicals  within  the  barrel  itself.  The  barrel  is 
rotated,  bringing  the  gold  intimately  into  contact  with  the  chlorine, 
which  acts  powerfully  upon  the  gold  while  in  the  nascent  condition. 
Barrel  chlorination  requires  a  large  initial  investment  in  machinery 
and  apparatus.  It  is,  however,  suited  to  a  large  tonnage,  and  is  a 
less  costly  process  and  gives  a  high  and  rapid  extraction. 

34.     THE  VAT  OR  PLATTNER  PROCESS  OF  CHLORINATION. 

The  ore,  which  as  above  stated,  has  been  subjected  to  an  oxidiz- 
ing roast  to  free  the  gold  and  render  the  ore  porous,  is  moistened 
and  charged  into  a  tank  or  vat  8  to  9  ft.  diam.  by  3  to  3.5  ft.  deep. 


Fig.    58.      CHLORINATION  LEACHING  VAT. 

This  vat,  Fig.  58,  has  a  false-bottom  of  perforated  1-in.  boards  sup- 
ported on  1-in.  strips  above  the  bottom  of  the  vat. 

Upon  the  perforated  bottom  is  spread  a  layer  of  quartz  first  in 
pieces  the  size  of  eggs,  then  smaller  toward  the  top.  Above  this  is  a 
2-in.  layer  of  sand,  and  on  this  is  laid  a  canvas  filter-cloth  or  an  open 


140 


THE    METALLURGY 


grating  of  boards,  to  protect  the  filter  and  give  a  surface  on  which 
to  shovel.  Chlorine  gas  is  admitted  through  the  pipe  n,  and  the  solu- 
tion is  discharged  through  the  hose  b,  into  the  launder  c,  which  leads 
to  the  settling  vats.  When  not  in  use  the  hose  is  turned  upward  in 
the  position  shown. 

The  4-ton  charge  of  ore  is  carefully  moistened  with  a  suitable 
amount  of  water  (6%).  If  too  dry,  the  gold  is  not  well  acted  upon 
by  the  chlorine ;  if  too  wret,  the  gas  does  not  penetrate  the  ore  in  a 
suitable  manner.  A  layer  of  dry  ore  is  first  scattered  over  the  bot- 
tom to  soak  up  the  water  left  in  the  filter,  that  the  gas  may  pass 
upward  freely.  The  charge  is  then  thrown  into  the  vat  through  a 
%-in.  screen,  which  breaks  the  lumps,  and  causes  the  ore  to  scatter 
loosely.  When  the  vat  has  been  filled  a  foot  deep,  the  gas  is  intro- 


Fig.    59.      CHLORINE   GENERATOR. 


duced  from  below  and  begins  to  rise  through  the  ore.  Charging  is 
continued  and  the  vat  is  filled  within  3  in.  of  the  top,  and  burlap- 
sacking  is  spread  over  the  surface.  The  cast-iron  cover  d,  is  then 
brought  by  an  overhead  crane  to  the  vat,  and  lowered  upon  it ;  and 
the  joint  between  the  cover  and  the  top  of  the  vat  is  made  tight  with 
clay-mortar  and  a  strip  of  cloth.  Chlorine  gas,  generated  in  a  sepa- 
rate vessel,  is  allowed  to  enter  the  vat  for  5  to  12  hours,  according 
to  the  fineness  of  the  gold.  The  finer  the  gold  the  faster  is  it  chlor- 
inated. The  charge  is  known  to  be  sufficiently  saturated  with  chlor- 
ine when,  upon  opening  the  stopper  e,  on  the  cover,  fumes  of  escaping 
chlorine  gas  can  be  detected  by  the  odor.  After  this  the  covered  vat 
is  allowed  to  stand  24  to  40  hours  to  chlorinate. 


OF   THE    COMMON    METALS.  141 

In  presence  of  moisture  the  soluble  tri-chloride  of  gold  is  formed 
as  follows : 

H2O  +  Au  +  3C1  =  AuCl3H,O. 

The  chlorine  is  produced  in  a  generating  vat  (Fig.  59).  This  is 
of  cast-iron,  lead-lined,  and  has  a  heavy,  tight  cylindrical  cover  c. 
For  a  4-ton  charge  of  ore  the  vessel  should  be  24  in.  diam.  by  12  in. 
deep.  It  is  charged  by  lifting  the  cover  and  putting  in  the  solid 
chemicals,  or  adding  these  through  the  plug  li.  The  chemicals  con- 
sist of  20  to  27  Ib.  dioxide  of  manganese,  27  to  32  Ib.  common  salt, 
and  40  to  60  Ib.  sulphuric  acid  of  66  °B.  The  acid  is  added  through 
the  funnel-tube  i,  and  is  followed  by  24  to  33  Ib.  water.  The  cover 
c,  has  a  water  seal  ;',  which  prevents  the  escape  of  the  gas.  The 
generator  stands  upon  a  sand-bath  q,  or  preferably  on  a  steam-bath 
or  coil,  by  which  it  is  heated  to  60°C.,  the  best  temperature  for  gen- 
erating the  chlorine.  The  reaction  is  as  follows : 

2NaCl  +  Mn02  +  2H2SO4  =  2C1  +  Na2S04  +  MnS04  +  2H20. 

The  stirrer  e,  is  turned  by  the  handle  g,  from  time  to  time.  This 
revolves  on  a  projecting  pivot  at  the  bottom  and  has  an  inverted  cup 
/,  soldered  to  the  shaft  that  forms  an  air-tight  water- joint  with  the 
corresponding  socket  filled  with  water  like  that  of  the  cover  at  /. 
The  plug  d,  serves  to  discharge  the  exhausted  contents  of  the  gen- 
erator. Through  the  delivery-pipe  A1,  and  the  horizontal  tube  I,  the 
gas  passes  to  the  wash-bottle  o,  where  hydrochloric  acid,  if  present, 
is  absorbed  by  the  water  contained  in  the  barrel  p.  The  gas,  thus 
washed,  passes  on  by  the  tube  ra  and  n  to  the  chlorinating  tank, 
Fig.  58. 

The  charge,  having  remained  in  the  vat  an  average  of  48  hours 
it  is  ready  for  leaching.  For  leaching,  the  cover  is  removed,  and 
water  from  a  hose  is  run  in,  and  evenly  distributed  over  the  ore. 
As  soon  as  the  ore  becomes  saturated  and  covered  with  water,  the 
solution  is  allowed  to  escape  through  the  hose  6,  which  is  lowered  for 
that  purpose.  The  level  of  the  water  is  maintained  at  the  top  by  a 
further  supply  until  the  escaping  solution,  on  being  tested,  gives  no 
reaction  for  gold.  Two  tons  of  water  per  ton  of  ore  is  used.  The 
tailing  or  leached  ore  often  contains  silver,  and  this  when  in  suffi- 
cient quantity,  can  be  recovered  later  by  hyposulphite-lixiviation. 
Thus  at  the  Sierra  Buttes  mine,  California,  this  tailing  is  leached 
48  hours  in  vats  3.5  ft.  diam.  by  5  ft.  high,  with  3%  solution  of  sodi- 
um-hyposulphite. The  silver,  precipitated  from  the  filtrate  by  means 
of  sodium-sulphide,  is  collected  and  sold. 

The  gold  solution  from  leaching  is  conducted  through  a  wooden 


142  THE    METALLURGY 

launder,  either  to  a  settling  vat  from  which,  after  settling  for  sev- 
eral hours,  it  passes  to  the  precipitating  vat,  or  through  a  filter-bag, 
and  then  at  once  to  the  precipitating  vat.  When  obtained  from  an 
8-oz.  ore,  it  may  contain  in  solution,  1%  of  various  base-metal  sul- 
phates, 0.9%  metal-chlorides,  0.02%  gold  chloride,  and  0.2%  free 
chlorine. 

The  precipitating  tank  is  6  ft.  diam.  by  3  ft.  high,  painted  (as 
also  are  the  other  tanks),  with  hot  asphalt,  to  which  has  been  added 
Portland  cement.  The  precipitant  for  gold  is  a  solution  of  ferrous 
sulphate,  prepared  at  the  works  by  dissolving  scrap-iron  in  sulphuric 
acid.  It  precipitates  the  gold  as  follows : 

2AuCl3  +  6FeS04  =  2Au  +  Fe2Cl(.  +  2Pe2(S04)3. 

To  save  time,  the  precipitant  is  added  with  the  first  of  the  gold 
solution  entering  the  tank  and  enough  more  is  stirred  into  the  solu- 
tion to  complete  the  precipitation.  The  vat  is  then  covered,  and  the 
gold  is  allowed  to  settle  12  hours,  or  much  longer.  The  clear  super- 
natant solution  is  then  run  off,  preferably  through  a  filter-press  (See 
Fig.  78),  and  the  filtrate  received  into  a  sawdust  filter  to  recover  any 
particles  of  gold  that  escaped  the  press.  Precipitation  with  ferrous 
sulphate  has  the  disadvantage  of  being  slow.  The  purple  color,  due 
to  the  presence  of  a  trace  of  gold,  can  be  detected  in  the  solution 
days  after  precipitation.  It  is  inferior  to  hydrogen  sulphide  in  this 
respect.  The  residue  in  the  tank  is  allowed  to  remain,  and  fresh 
solution  in  turn  is  run  in  and  precipitated.  The  precipitate  thus 
gradually  accumulates.  Finally,  the  precipitating  tank  is  drained 
completely  through  the  filter-press,  and  the  residue  is  washed  with 
dilute  sulphuric  acid  to  remove  the  ferric  salt  remaining.  The  moist 
product  from  the  press  is  mixed  with  soda,  borax,  and  nitre,  and 
melted  in  graphite  crucibles.  The  molten  gold  is  poured  into  molds, 
and  after  removing  the  slag,  is  re-melted  with  a  little  borax  to 
obtain  a  uniform  bar.  The  gold  thus  obtained  is  920  to  990  fine,  the 
impurities  being  lead  and  iron.  The  extraction  or  recovery  of  the 
gold  is  90  to  92  per  cent. 

Cost  of  plant  and  treatment. — The  cost  of  erecting  a  plant  in 
California,  capable  of  treating  6  tons  daily,  is  $6000  to  $7000.  In 
1886  the  cost  of  chlorination  at  the  Providence  mine,  exclusive  of 
supervision,  interest,  and  depreciation,  was  $6.30  per  ton.  About 
one-half  of  this  ($3),  was  the  cost  of  roasting.  This  cost  has  since 
been  considerably  reduced  by  the  use  of  oil  fuel  and  by  the  em- 
ployment of  mechanical  furnaces  of  the  Edwards  type  (See  Fig.  39). 


OF    THE    COMMON    METALS. 


143 


35.     BARREL  CHLORINATION. 

An  example  of  an  ore  generally  treated  by  chlorination  is  that  of 
Cripple  Creek:   the  ore  (2)  of  the  table.     The  ore  contains  gold  tel- 


Fig.    60.      SECTION  THROUGH  CRUSHING  MILL  AND   ROASTER. 

luride,  and  must  be  roasted  to  release  the  gold  from  combination 
with  tellurium,  and  to  expel  all  sulphur  above  0.1%.  The  complete 
process  of  barrel-ohlorination  is  as  follows : 


144 


THE    METALLURGY 


The  ore  is  coarsely  crushed  in  breakers  and  rolls  to  %-in.  size 
or  less  and  placed  in  storage-bins.  The  crushing  is  done  during  the 
day-shift.  Thus  an  accident  to  either  the  coarse  or  the  fine-crushing 
system  need  not  stop  the  operation  of  the  mill. 

The  ore  from  any  of  the  storage-bins  is  drawn  off  as  needed  to 
the  feed-hopper  of  the  dryer.  It  is  dried  and  fine-crushed  precisely 
as  described  in  section  14,  except  that  it  need  not  be  crushed  finer 
than  10  to  16-mesh.  By  avoiding  fine  crushing,  less  dust  is  made, 
and  after  roasting,  the  product  is  pervious  to  the  attack  of  the 
chlorine  gas,  and  is  more  readily  leached. 

Cripple  Creek  ore  is  roasted  in  a  mechanical  furnace,  such  as 


Fig.61.      CHLORINATION   BARREL. 

the  Wethey,  Fig.  36  and  37,  or  the  Edwards,  Fig.  38  and  39.  The 
finishing  temperature  should  not  be  higher  than  necessary  to  break 
up  the  sulphates  formed  in  roasting.  In  operating  the  Wethey  furn- 
ace, the  lower  hearth  is  used  to  cool  the  roasted  ore.  The  cooling  is 
effected  in  the  Edwards  furnace  by  surrounding  the  troughs  of  the 
discharge  conveyor  by  water  pipes.  Ore  thus  cooled  is  raised  by  an 
elevator  to  storage-bins,  whence  it  is  drawn  as  needed  to  the  chlorina- 
tion  barrels. 

The  ehlorination  barrel. — These  barrels,  as  shown  in  perspective 
in  Fig.  61,  and  in  transverse  and  longitudinal  section  in  Fig.  62,  are 
rotated  on  a  horizontal  axis.  They  are  supported  on  trunnions,  and 
driven  at  12  rev.  per  min.  The  shell  d,  is  of  sheet-steel  with  heavy 


146  THE    METALLURGY 

cast-iron  ends  c,  and  provided  with  two  charging1  doors  or  man- 
holes 6,6.  The  cylinder  is  lined  with  sheet  lead  %  in.  thick,  bolted 
to  the  shell.  It  contains,  as  shown,  a  filter-frame,  or  diaphragm  //, 
of  hard  wood,  intended  for  filtering  the  clear  solution  after  treat- 
ment with  chlorine,  leaving  the  exhausted  ore  behind  in  the  barrel. 
The  filter  is  made  of  blocks  /  that  sustain  a  floor  of  lead  plates  %  in., 
thick,  perforated  with  %-in.  holes.  The  plate  itself  is  corrugated 
to  allow  circulation  of  the  filtrate  over  it.  Instead  of  these  plates, 
a  perforated  2-in.  plank  floor  has  been  used.  Upon  the  plate  (or 
floor),  rests  a  lead  sheet  of  4  Ib.  per  sq.  ft.  This  sheet  is  perforated 
with  0.05-in.  holes,  %  in.  between  centers.  To  hold  down  the  filter- 
sheet,  a  wooden  frame  or  grating  is  placed  upon  it.  This  is  held  by 
blocks  //,  and  heavy  strips  i,  securely  bolted  to  the  barrel.  The  wood 
frames  beneath  last  three  months;  those  above,  but  two  or  three 
weeks.  This  wood-work,  if  immersed  in  boiling  tar  or  asphalt  until 
thoroughly  impregnated,  lasts  longer  and  absorbs  but  little  solution. 
In  place  of  lead  filters,  a  woven  asbestos  cloth  may  be  used.  The 
cloth  needs  renewal  after  50  to  60  charges  have  been  treated.  Bar- 
rels have  been  made  6  ft.  diam.  by  16  ft.  long,  with  a  capacity  of 
18  tons.  The  common  size,  however,  is  6  ft.  diam.  by  12  ft.  long, 
and  the  capacity  is  8.5  tons. 

Charging  the  ore. — Crushed,  roasted,  and  cooled,  the  ore,  in 
weighed  charges,  is  conveyed  from  the  storage-bins  in  two-wheeled 
buggies,  and  placed  in  the  charging  hoppers  that  belong  to  each 
cylinder  (See  the  charging  hoppers  of  the  Bruckner  roasting-f urn- 
ace,  Fig.  35).  Into  the  cylinder  is  first  run  80  to  140  gal.  water  per 
ton  of  ore,  enough  to  make  with  the  ore,  an  easily  flowing  pulp. 
Next,  a  measured  quantity  of  sulphuric  acid  is  added,  and  then  the 
charge  of  ore.  Finally  a  weighed  amount  of  'bleach'  or  bleaching 
powder  (CaCl2O)  is  added.  The  quantity  of  chemical  to  be  added 
to  the  charge  depends  on  the  nature  of  the  ore,  and  is  determined  by 
experiment.  On  roasted-  Cripple  Creek  ore  12  to  15  Ib.  bleaching 
powder  of  34  to  36%  available  chlorine,  and  24  to  30  Ib.  sulphuric 
acid  of  66°B.  is  used  per  ton  of  ore. 

The  charge-openings  of  the  barrel  are  now  closed.  The  barrel  is 
started  to  slowly  revolving  (12  rev.  per  min.),  for  a  period  of  1  to  4 
hours.  In  the  case  of  Cripple  Creek  ore,  for  3  hours.  The  chemicals 
enter  into  thorough  contact  with  one  another  and  with  the  ore,  and 
react  as  follows : 

(1)     2CaOCl2  +  2H2S04  ==  4C1  +  2CaSO4  +  2H20. 
Chlorine,  in  nascent  condition,  acts  with  greater  energy  upon  the 


OF    THE    COMMON    METALS.  147 

gold  then  chlorine  formed  outside  the  barrel.    With  the  gold  it  forms 
a  soluble  gold  chloride  thus : 

(2)     Au  +  3C1  +  H2O  ==  AuCl,,.H20. 

To  determine  when  the  ore  is  thoroughly  saturated  with  chlo- 
rine, a  stop-cock  j  is  opened  and  gas  that  issues  is  tested  for  chlorine 
with  ammonia  which  produces  a  white  fume  of  NH4C1  with  chlorine. 
If  free  chlorine  is  not  detected,  the  barrel  is  stopped,  opened,  and 
more  bleach  and  acid  added.  In  place  of  using  chloride  of  lime  and 
sulphuric  acid,  as  above  described,  chlorine  (about  1  Ib.  per  ton  of 
ore),  has  been  introduced  into  the  barrel  in  liquid  form.  Chlorine 
can  be  obtained  in  this  form,  in  strong  steel  cylinders  or  drums. 

After  saturation  with  chlorine  the  barrel  is  revolved  an  hour, 
then  stopped  in  position  for  filtering  with  the  filtering  diaphragm 
down  and  level.  The  outlet  pipe  K,  is  connected  by  a  hose  to  the 
settling  tank  and  opened ;  and  water  is  pumped  into  the  barrel  above 
the  charge  through  the  valve  /.  The  solution  is  now  drained  off, 
and  the  water  above  the  charge  forced  rapidly  through  by  means  of 
compressed  air  introduced  through'  the  valve  j.  The  excess  of  chlo- 
rine is  absorbed  by  the  wash-water,  and  does  not  enter  the  building. 
The  operation  of  filtering  is  suspended,  connections  are  broken, 
valves  closed,  and  the  barrel  revolved  a  few  times  to  mix  the  con- 
tents again,  and  to  break  up  channels  that  may  have  formed  during 
the  leaching.  The  barrel  is  then  stopped,  water  run  in,  compressed 
air  admitted,  and  the  washing  resumed.  This  is  repeated  until  no 
gold  is  found  in  the  escaping  filtrate  when  tested.  The  compressed 
air  admitted  is  under  a  pressure  of  40  Ib.  per  sq.  in.  The  time  of 
filtering  and  washing  on  an  average  is  2%  hours.  The  water  used 
is  50%  the  weight  of  the  ore.  All  connections  are  finally  broken, 
valves  closed,  and  man-holes  opened,  and  the  cylinder  is  revolved 
several  times  to  discharge  the  contents.  It  is  then  washed  out  with 
a  hose  to  prepare  it  for  another  charge. 

The  solution  from  the  barrels  runs  into  lead-lined  or  asphalt- 
coated  settling-tanks  10  ft.  diam.  by  7%  ft.  high.  Here  any  sediment 
which  escaped  the  filter  is  settled  in  about  8  hours.  The  clear  sup- 
ernatant solution  is  withdrawn  at  a  point  10  in.  above  the  bottom  of 
the  settling-tank  to  avoid  disturbing  the  sediment,  and  run  into  the 
gold-solution  tanks  on  a  lower  level,  where  it  is  stored.  The  clear 
solution  from  the  gold-solution  tank  is  pumped  through  the  opening 
.4,  Fig.  63,  into  the  precipitation  tank  X,  which  is  10  ft.  diam.  by  12 
ft.  high.  The  free  chlorine  is  removed  by  passing  sulphur  dioxide 
through  the  solution  from  the  SO2  generator  //.  This  generator 


148 


THE    METALLURGY 


is  a  cast-iron  receptacle  containing  a  pan  F  in  which  sulphur  is 
burned.  Compressed  air,  admitted  from  a  pipe  W,  supplies  oxygen 
to  burn  the  sulphur  to  S02  and  pressure  to  drive  the  fumes  into  the 
solution  along  the  pipes  o,v  and  the  lead  pipe  r,  through  numerous 


v- 


Fig.    63.      PRECIPITATION  APPARATUS  OF   BARREL  CHLORINATION. 

small  holes  where  the  horizontal  part  of  the  pipe  crosses  the  bottom 

of  the  tank.    The  sulphur  dioxide  acts  upon  the  chlorine  as  follows: 

(3)     C12  +  SO2  +  2H2O  ==  H,SO4  +  2HC1. 

Sulphuric  and  hydrochloric  acids  are  formed  by  the  action  of 
sulphurous  acid  upon  chlorine,  and  both  remain  in  aqueous  solution. 
The  operation  is  completed  in  15  to  20  minutes  as  indicated  by  test- 
ing the  solution  with  H2S,  when  a  permanent  precipitate  forms. 

The  chlorine  having  been  removed,  we  are  ready  to  precipitate 
the  gold  by  passing  in  H2S.  To  do  this,  the  pipe  v  is  connected  to  the 
lead-lined  generator  G  which  contains  lumps  of  iron  sulphide  resting 


OF    THE    COMMON    METALS.  149 

on  a  perforated  lead  false-bottom.     Dilute  sulphuric  acid  admitted 
below  the  false-bottom  comes  in  contact  with  the  iron  sulphide  and 
abundantly  generates  H2S  according  to  the  equation  : 
(4)     FeS  +  H2SO4  =*  FeS04  +  H2S. 

Compressed  air,  entering  by  the  pipe  w  and  the  valve  c,  drives  the 
gas  through  the  pipe  v  and  the  lead  pipe  r  into  the  solution  in  the 
tank,  and  precipitates  the  gold  as  follows : 

(5)     2AuCl3  +  3H2S  =  Au2S3  +  6HC1. 

The  gold  is  thus  thrown  down  as  an  auric  sulphide  in  a  solution 
containing  both  sulphuric  and  hydrochloric  acids.  The  precipitation 
is  rapid,  taking  about  10  minutes,  and,  it  is  possible  by  selective  pre- 
cipitation and  careful  working  to  leave  copper  in  solution  while  pre- 
cipitating the  gold.  Where  the  chemicals  in  the  generator  become 
exhausted,  the  waste  liquid  is  discharged  into  the  waste-launder  p. 
The  chemicals  consumed  per  ton  of  roasted  ore  are,  1  Ib.  iron  sulph- 
ide, ^4  MX  sulphur,  and  2%  Ib.  sulphuric  acid. 

More  recent  treatment  does  not  include  the  use  of  S02  gas,  H2S 
gas  alone  being  used.  At  first,  H2S  is  oxidized  by  the  chlorine,  thus: 

(6)     H2S  +  8C1  +  4H20  =  H2S04  +  8HC1. 

Sulphuric  and  hydrochloric  acids  are  formed  by  this  reaction  after 
which  auric  sulphide  is  precipitated  as  in  equation  (4). 

After  being  precipitated,  Au2S3  is  allowed  to  settle  two  hours. 
The  clear  solution  then  is  drawn  off  at  C,  10  in.  above  the  bottom  of 
the  tank,  through  the  pipe  n  into  the  filter-press,  where  it  is  filtered 
under  its  own  hydrostatic  head  of  25  ft.  This  is  done  to  recover  any 
possible  flakes  of  gold  sulphide  that  failed  to  settle  in  the  tank  x.  In 
three  or  four  hours  after  precipitation,  the  tank  can  receive  a  fresh 
charge  of  gold-bearing  solution. 

The  precipitate  collects  upon  the  bottom  of  the  tank,  and  after 
several  charges  have  been  treated,  the  united  precipitate  is  drawn 
off  at  D  and  delivered  through  the  man-hole  L  into  the  pressure-tank 
2,  by  the  hose  y.  The  precipitating  tank  is  then  washed  clean  with 
the  aid  of  a  hose.  The  pressure-tank  is  4  ft.  diam.  by  4%  ft.  high. 
When  charged,  the  cover  L  is  clamped  in  place,  and  compressed  air, 
under  a  pressure  of  40  Ib.  per  sq.  in.,  is  admitted  through  t.  At  the 
same  time  connection  is  made  to  the  filter-press  T  through  the  pipe 
u,  and  the  precipitate  collects  in  the  press  under  the  above  pressure. 
This  filter-press  is  more  clearly  illustrated  in  Fig.  78.  A  set  of  filter- 
frames  outlasts  the  treatment  of  6000  to  8000  tons  of  ore.  The  filt- 
rate from  the  press  passes  over  a  sawdust  filter-bed,  as  a  safeguard, 
before  it  is  run  to  waste.  The  sawdust  is  collected  occassionally,  and 


150  THE    METALLURGY 

burned,  to  recover  the  small  amount  of  gold  which  it  may  have 
caught. 

The  precipitate  of  gold  sulphide  also  contains  sulphur,  and  sul- 
phides of  arsenic,  antimony,  copper,  and  silver,  forming  a  'sulphide 
cake'.  The  press  is  next  opened  and  the  cake  is  withdrawn.  While 
still  moist  it  is  placed  in  trays  44  in.  long,  24  in.  wide,  4  in.  deep, 
and  mixed  with  borax,  nitre,  and  soda.  The  trays  are  then  placed 
in  cast-iron  muffles  that  are  heated  by  coal,  and  connected  to  a  flue- 
chamber,  where  any  mechanically  escaping  flue-dust  is  caught.  The 
precipitate  dries  here,  and  after  drying,  the  temperature  is  raised  to 
decompose  and  expel  the  sulphides,  the  operation  taking  an  hour. 
The  roasted  material  has  a  light-brown  color,  and  contains  70  to 
80%  of  gold.  It  is  carefully  transferred  to  a  crucible  and  melted  in 
a  wind-furnace.  The  content  of  the  crucible,  including  the  slag,  is 
poured  into  a  conical  mold  where  the  gold  collects  at  the  bottom. 
Upon  cooling,  the  gold  ingot,  900  to  950  fine,  is  separated  from  the 
slag,  remelted,  and  cast  into  a  bar  suitable  to  be  forwarded  to  the 
United  States  mint.  After  deducting  a  charge  of  2c.  per  ounce  for 
melting  and  assaying,  the  gold  at  the  mint  should  net  $20.65  per 
ounce  of  contained  gold. 

The  slag,  from  the  melting  operation,  may  be  melted  again  with 
one-seventh  its  weight  of  litharge  with  a  reducing  agent,  and  a  lead 
button  obtained.  This  lead  can  be  scorified  several  times,  until  it 
is  reduced  to  a  small  button,  and  finally  cupelled  to  recover  the  small 
amount  of  additional  gold. 

The  cost  per  ton  of  treating  Cripple  Creek  ores,  on  a  large  scale, 
by  barrel  chlorination,  is  as  follows : 

Labor,   including  salaries    $1.34 

Chemicals  and  supplies   0.72 

Fuel,  roasting  and  power 0.70 

Renewals  and  repairs    0.45 

General   expense    0.32 


Total  cost  per  ton  $3.53 

The  cost  of  sulphuric  acid,  66°B.,  is  0.9  to  l.lc.  per  Ib. ;  bleaching 
powder  1.8c.  (New  York)  ;  sulphide  of  iron  3c.  (this  can  be  made  at 
the  works  from  scrap  wrought-iron  and  sulphur)  ;  sulphur  2  cents. 
The  water  needed  is  approximately  two  tons  per  ton  of  ore  treated, 
which  includes  that  used  for  power.  This  quantity  can  be  reduced 
if  settling  tanks  are  employed  and  the  water  used  again. 


OF    THE    COMMON    METALS. 


151 


Following  is  a  flow-sheet  indicating  the  procedure  in  the  barrel- 
chlorination  process  above  described. 


MlNEORE 
I 


CRUSHING    PLANT 
vSft  FIG  2£) 


CRUSHED  RAW  ORE 

^ 

|  STORAGE  BINS     [ 


|  BELT  ELEVATOR) 

3 

EXTRACTION  - 

^ 

(FEED   HOPPER  | 

X 

[ROASTING  FURNACE  | 

RO/\ST!D  CRE 

^ 

[COOLING  HEARTH&  FLOOR] 

X 

ROASTED  Cooueo  ORE 


[STORAGE   SINS 


FLUX  e 


INGOTS 

-To    MARKET 


-SOLUTION 


Fig.    64.      FLOW-SHEET   OF   BARREL   CHLORINATION. 

36.     CYANIDATION  OF  GOLD  (AND  SILVER)  ORES. 

The  process  of  cyanidation  consists  in  attacking  the  gold  or  silver 
contained  in  ores  with  dilute  solutions  containing  0.5  to  0.7%  of 
potassium  cyanide.  If  the  ore  contains  pyrite  it  may  be  acid  as  the 
result  of  weathering.  Before  applying  potassium  cyanide  this  acid- 
ity must  be  corrected  by  adding  caustic  soda  or  quick-lime.  The 
gold-bearing  cyanide  solution  is  filtered  off,  and  the  precious  metals 
contained  are  precipitated  upon  zinc  shavings,  zinc  dust,  or  by  elec- 
trolysis. The  precipitate  is  collected,  treated  to  purify  it,  melted, 
and  cast  into  bars  or  ingots  of  metal.  The  solution,  free  from  gold, 
called  'barren  solution',  is  used  again. 

The  cyanide  precess  is  a  success  with  many  ores.  The  field  of  its 
usefulness  is  extending,  not  only  in  the  variety  of  ores,  but  ores  con- 
taining smaller  and  smaller  amounts  of  metal  are  being  treated.  An 
important  advantage  of  cyanidation  over  chlorination  is  that  roast- 
ing by  no  means  is  essential  though  sulphides  be  present.  Sulphides 
containing  silver,  however,  must  be  small  in  quantity,  if  silver  is  to 
be  extracted  in  a  period  of  time  compatible  with  economy. 

Dilute  cyanide  solution  has  a  selective  action  in  extraction  and 
can  dissolve  gold  and  silver  from  an  ore,  at  the  same  time  attacking 
but  little  the  base-metals  that  are  present.  If,  however,  copper  is 
present  in  soluble  form  it  consumes  cyanide,  though  in  modern  prac- 
tice a  small  quantity  is  not  considered  a  serious  obstacle. 

Ores  suited  to  cyanidation. — With  reference  chiefly  to  gold,  we 
have  the  following  classes  of  ores : 


152  THE    METALLURGY 

(1)  Free-milling  ore. — Those  in  which  gold  occurs  in  a  fine  or 
microscopic  form,  and  in  which  panning  reveals  to  the  naked  eye 
few  or  no  visible  colors  of  gold.    Where  gold  particles  are  coarse  the 
time  required  to  dissolve  them  is  long,  and  cyanidation  becomes  im- 
practicable.    It  is  possible,  however,  to  remove  coarse  gold  by  am- 
algamation,, and  after  so  doing  the  residue  can  be  treated  by  cyanida- 
tion. 

(2)  Telluride  ores. — Before  these  ores  are  cyanided  they  must 
be  treated  to  liberate  the  gold  from  its  combination  with  tellurium. 
When  rich  and  containing  telluride  in  spots,  roasting  develops  shots 
or  particles  of  gold,  and  these  must  be  removed  by  amalgamation  or 
concentration  before  cyaniding. 

(3)  Pyrite  ores. — The  gold  in  such  ores  is  supposed  to  occur  in 
the  form  of  films  on  the  surface  of  the  pyrite  crystals.    When  these 
are  crushed  fine  to  expose  fresh  faces,  the  gold  can  be  dissolved  by 
cyanide.     Some  pyrite  ores,  however,  contain  the  gold  within  the 
substance  of  the  crystals,   and  these  ores  must  be  roasted  before 
cyaniding.    Eoasting  improves  ore  by  rendering  it  more  porous,  and 
hence  more   accessible  to  the  solution.     It  destroys  the  colloid  of 
slime,  and  makes  a  material  more  easily  filtered. 

(4)  Talcose  or  clayey  ores. — These,  when  crushed  for  cyaniding, 
produce  much  slime  that  is  impenetrable  by  solution,  and  can  not  be 
leached.     Slime  is  treated  in  a  way  that  adds  to  the  cost  of  extrac- 
tion.    Roasting,  as  stated,  would  accomplish  this  end,  but  the  cost 
would  be  prohibitory  in  some  cases. 

37.  DEVELOPMENT  OF  THE  CYANIDE  PROCESS. 

This  process,  originally  called  the  MacArthur-Forrest  process, 
was  patented  and  vigorously  advanced  by  the  inventors.  It  was  first 
put  in  practice  at  the  Robinson  mine  on  the  Rand,  South  Africa,  and 
applied  to  the  recovery  of  gold  in  large  tailing-dumps,  that  had  ac- 
cumulated in  stamp-milling  and  had  been  impounded  behind  crib- 
work,  or  low  dams.  This  tailing  was  leached  in  filter-bottom  vats 
with  cyanide  solutions.  The  gold-bearing  filtrate  was  passed  through 
zinc-boxes,  containing  zinc  shavings,  to  precipitate  the  gold,  and 
the  precipitate  was  purified  and  the  gold  collected  and  cast  in  ingots. 
So  long  as  these  dumps  lasted  the  simple  process  was  sufficient,  but 
when  it  became  necessary  to  treat  the  tailing  immediately  from  the 
mill,  another  kind  of  practice  was  begun.  This  'second  method'  in 
outline  is  as  follows:  The  ore  is  crushed  and  amalgamated  in  the 
customary  way.  The  tailing  from  the  amalgamated  plate  is  classi- 


OF    THE    COMMON    METALS.  153 

fied  into  two  products,  one  of  which  consists  of  coarse  sand,  the  other 
of  slime  and  fine  sand.  The  coarse  sand  is  more  readily  leached  when 
freed  from  slime.  To  correct  the  acidity,  milk  of  lime  is  added  to  the 
finer  portion  as  it  passes  through  the  launder  to  the  leaching-vats. 
The  sand  here  settles  and  the  slime  passes  on,  suspended  in  the  escap- 
ing water.  When  the  tank  is  filled  it  is  drained,  and  the  contained 
water  is  replaced  by  weak  cyanide  solution.  Discharge-valves  or 
doors  (See  Fig.  68),  are  opened  and  the  contents  of  the  vat  are 
shoveled  into  a  vat  directly  beneath.  A  solution  of  0.15  to  0.25% 
cyanide  is  allowed  to  percolate  some  days  through  the  sand,  and  then 
is  displaced  by  a  weaker  solution,  and  the  weaker  one  by  water.  The 
gold  in  the  gold-bearing  filtrate  is  precipitated  on  zinc  shaving,  and 
the  solution,  after  removing  the  gold,  is  used  again.  This  is  called 
'double  treatment',  since  the  extraction  of  gold,  begun  in  one  vat. 
is  completed  in  a  second  one.  While  a  little  more  expensive  in  equip- 
ment and  in  handling,  it  is  more  thorough.  The  coarse  sand  requires 
much  longer  time  to  leach  than  the  fine  sand,  and  hence  is  treated 
separately. 

The  slime  is  caught  in  large  tanks  and  allowed  to  settle.  After 
settling,  the  clear  supernatant  water  is  drawn  off.  A  0.01  to  0.02% 
solution  of  cyanide  is  added  to  the  thickened  slime  remaining  in  the 
tank,  and  the  whole  is  agitated  by  a  mechanical  stirrer  for  some 
hours.  It  is  then  settled,  and  the  clear  gold-bearing  solution  is  drawn 
into  a  gold-solution  tank,  from  which  it  is  supplied  in  a  regulated 
stream  to  the  zinc-boxes  for  precipitation.  The  residual  slime  is 
treated  again  with  weak  solution,  settled,  and  decanted.  The  spent 
slime  then  is  washed  out  to  the  dump. 

This  method  of  treatment  is  slow,  and  requires  a  large  number  oi 
tanks ;  and  gold  is  carried  away  in  the  solution  that  remains  in  the 
slime.  To  overcome  these  difficulties  filter-pressing  has  been  tried. 
The  slime,  after  agitating  as  above  described,  with  cyanide  solution, 
is  pumped  through  a  filter-press  (Fig.  78),  where  it  is  separated  and 
washed,  and  the  filtrate  is  collected  in  the  gold-solution  tank,  while 
the  exhausted  slime  removed  from  the  press  is  thrown  away.  The 
trouble  with  filter-pressing,  as  ordinarily  practiced,  is  that  it  is  ex- 
pensive and  slow.  A  number  of  costly  presses  are  required  for  a 
large  plant  and  these  consume  power  and  require  labor  to  operate 
them.  In  the  United  States  this  has  led  to  the  adoption  of  suction- 
filters  of  the  Moore  type.  In  South  Africa,  Western  Australia,  and 
elsewhere,  the  Butters  and  Ridgway  filters  have  given  good  service, 
while  pressure-tank  filters  of  the  Blaisdell  type,  that  use  filter-leaves 
in  the  name  general  way*  have  advocates. 


154  THE    METALLURGY 

Western  Australian  ores  require  fine  grinding,  to  liberate  the 
gold.  While  fine  material  cannot  be  leached  in  vats,  it  can  be  filter- 
pressed.  This  has  led  to  the  use  of  the  grinding  pan  as  in  silver 
milling,  and  to  the  tube-mill  in  particular,  which  has  proved  particu- 
larly well  adopted  to  fine  grinding.  The  outline  of  the  process  as 
practised  in  Western  Australia  is  as  follows : 

The  ore  is  coarsely  crushed  in  rock-breakers,  then  dry-crushed 
either  in  a  Griffin  mill,  a  ball  mill,  or  rolls.  Western  Australian  ore 
containing  tellurides  must  be  roasted  to  decompose  and  expel  the 
tellurium,  and  liberate  the  gold  for  attack  by  cyanide  solution.  By 
roasting,  the  ore  becomes  porous,  and  colloidal  compounds  are  de- 
stroyed, yet  it  is  necessary  to  grind  fine  if  a  high  extraction  is  to  be 
obtained.  The  pulverizing  has  been  done  in  grinding-pans  or  tube- 
mills,  in  the  presence  of  cyanide  solution.  The  product  is  sent  to  a 
settler  to  remove  large  particles  of  gold,  if  present,  and  all  the  pulp 
is  then  filter-pressed,  and  washed  in  the  press,  the  process  giving  an 
extraction  of  93  per  cent. 

The  trend  of  modern  practice  is  toward  fine  grinding,  and  treat- 
ment of  the  whole  product  as  slime,  under  the  idea  that  the  fine  re- 
grinding  of  the  product  insures  pulverizing  the  coarse  gold,  so  that 
it  can  be  cyanided  in  reasonable  time.  The  alternative  process  to 
this  is  to  finely  grind  the  product,  classify  it  into  sand  and  slime,  and 
treat  the  sand  by  percolation  and  the  slime  by  pressure  of  suction 
filtration. 

38.     THE  CHEMISTRY  OF  THE  CYANIDE  PROCESS. 

When  a  solution  containing  0.5%  potassum  cyanide  is  brought 
into  contact  with  finely-ground  ore  containing  gold  in  microscopic 
particles,  the  gold  gradually  goes  into  solution.  The  reaction  that 
occurs  was  first  shown  by  Eisner,  and  is  called  Eisner's  equation.  It 
is  as  follows : 

2Au  +  4KCN  +  O  +  H20  =  2AuK(CN)2  +  2KOH 
The  gold  is  dissolved  by  the  action  of  potassium  cyanide  in  the  pres- 
ence of  oxygen  and  water;  the  compounds  formed  are  auro-potas- 
sium  cyanide  and  caustic  potash. 

Silver,  but  more  slowly,  dissolves  according  to  a  similar  reaction, 
thus : 

2Ag  +  4KCN  +  0  +  H20  =  2AgK(CN)2  +  2KOH.     . 

It  is  noticed  that  oxygen  is  needed  to  fulfil  the  requirements  of 
the  reaction,  and  consequently  ore,  or  solution  acting  on  ore,  must  be 
in  some  manner  aerated.  When  oxygen  of  the  dissolved  air  is  con- 
sumed, action  ceases,  but  it  proceeds  again  with  access  of  fresh  air. 


OF    THE    COMMON    METALS.  155 

Oxidizing  agents  like  potassium^  chloride-  and  permanganate,  and 
the  peroxides  of  lead,  manganese,  and  sodium  may  be  used  to  furnish 
oxygen,  in  place  of  air,  but  are  too  expensive  for  practical  use. 

When  an  ore  containing  pyrite  is  exposed  to  the  weather,  air  and 
moisture  slowly  act  upon  the  pyrite,  according  to  the  reaction : 

3FeS2  +  2H2O  +  22O  =  FeS04  +  Fe2(SO4)3  +  2H2SO4 
Ferrous  and  ferric  sulphates  and  sulphuric  acid  are  thus  formed. 
The  two  first  named  would  tend  to  precipitate  gold.  Ferric  sul- 
phate is  acid  in  its  reaction,  and  with  the  sulphuric  acid,  if  un- 
neutralizecl  it  would  decompose  and  cause  a  serious  loss  of  potassium 
cyanide.  Such  compounds  are  called  'cyanicides'.  Ore,  therefore, 
that  contains  pyrite,  and  that  has  lain  exposed  to  the  weather,  needs 
caustic  soda  or  quick-lime  Ca(OH)2  to  neutralize  the  acidity,  and  an 
excess  to  provide  for  the  acid  resulting  from  further  decomposition. 
This  excess  is  the  'protective  alkalinity'  as  it  is  called.  To  remove 
the  soluble  ferrous  sulphate  and  sulphuric  acid,  a  water-wash  before 
treatment  would  be  sufficient.  This  would  require  much  time  and  in 
place  of  it  quick-lime  is  added.  Quick-lime  acts  upon  the  products 
of  the  above  reaction  as  follows : 

FeSO4  +  Ca(OH)2  =  Fe(OH)^+  CaS04 
Fe2(SO4)3  +  3Ca(OH)2  =  2Fe(OH)3  +  3CaS04 

H2S04  +  Ca(OH)2  =  2H20  +  CaS04 

The  reaction  causes  the  formation  of  a  harmless  iron  hydroxide  and 
calcium  sulphate. 

The  reaction  that  occurs  when  gold-bearing  solution  comes  in 
contact  with  zinc  shaving  in  the  precipitating  boxes,  or  when  'zinc 
dust'  is  stirred  into  such  a  solution,  is  generally  assumed  to  be  a 
simple  replacement  of  zinc  by  gold,  thus : 

2KAu(CN)2  +  Zn  =  K2Zn(CN)4  +  2Au 

The  gold  falls  as  a  brown  or  black  precipitate,  while  the  zinc  potas- 
sium cyanide  remains  in  solution. 

The  barren  cyanide  solution  from  which  gold  has  been  removed 
is  used  repeatedly,  and  accumulates  impurity  from  the  ore,  and  from 
the  zinc  with  which  it  was  in  contact  in  the  zinc-boxes.  In  conse- 
quence it  becomes  gradually  less  efficient  than  fresh  solution.  It  has 
been  found  that  the  addition  of  quick-lime  increases  the  solvent  ac- 
tion of  cyanide  solution  upon  pure  quartz  ore,  but  it  is  without  effect 
on  ore  containing  sulphide.  Such  solution,  if  treated  with  sodium 
sulphide  to  the  point  of  exact  neutrality,  and  with  a  small  excess  of 
lead  acetate,  and  given  time  to  permit  the  resultant  sulphide  to  pre- 
cipitate, is  improved  in  extractive  power  thus : 


156  THE    METALLURGY 

K2Zn(CN)4  -f  Na2S  =  K2Na2(CN)4  +  ZnS 

The  cyanide  is  here  regenerated,  while  the  zinc  sulphide  separates. 
This  is  a  means  of  overcoming  the  accumulation  of  zinc  in  solution 
which  is  one  of  the  drawbacks  to  the  use  of  zinc  for  precipitation 
compared  with  electrical  deposition. 

We  are  not,  however,  altogether  dependent  upon  chemicals  to  dis- 
pose of  the  zinc.  We  have  shown  by  Eisner's  equation  that  caustic 
potash  is  set  free.  Caustic  potash  reacts  upon  sulphide  contained  in 
the  ore  and  causes  the  formation  of  a  soluble  sulphide  which  in  turn 
reacts  like  the  sodium  sulphide  in  the  reaction  above,  and  precipi- 
tates zinc  sulphide  upon  the  ore. 

The  following  minerals  and  chemical  compounds  destroy  or  com- 
bine with  cyanide,  and  render  it  incapable  of  dissolving  gold : 

Copper  in  the  form  of  sulphate,  carbonate,  copper  glance,  erubes- 
cite,  or  copper  pyrite.  The  sulph-antimonites  of  copper  are  without 
action.  Manganese  as  'wad'  (impure  hydrous  oxide),  but  not  the 
carbonate  or  oxide.  Zinc  as  calamine,  but  not  blende  or  zinc  silicate. 
A  distinct  loss  is  traceable  to  the  presence  of  organic  matter  like 
grass-roots,  decayed  wood,  etc. 

To  increase  the  activity  of  zinc  shaving  in  precipitating  of  gold 
it  may  be  dipped  in  a  10%  solution  of  lead  acetate  just  before  it  is 
placed  in  the  zinc-boxes.  Lead  is  precipitated  upon  the  surface  of 
the  shavings,  forming  zinc-lead  couples,  which  electrically  react  upon 
the  gold  in  solution. 

39.     CLASSIFICATION  OF  CYANIDATION  METHODS. 

First  method. — For  oxidized  ore  which  forms  but  a  moderate 
amount  of  slime  when  crushed  and  can  be  leached,  and  in  which  the 
contained  gold  is  fine  and  free.  Example :  Mercur,  Utah. 

Second  method. — For  ore  that  crushes  with  the  formation  of  slime 
and  contains  both  coarse  and  fine  gold  (coarse  gold  can  be  recovered 
by  amalgamation),  and  pyrite  and  other  sulphides,  that  can  be  re- 
moved by  concentration.  The  tailing  is  separated  into  sand  and 
slime.  The  sand  is  treated  by  leaching.  The  slime  is  either  rejected 
or  treated  by:  (a)  Decantation. — Best  example:  South  Africa,  (b) 
Filter-pressing. — Best  example :  Homestead  mine,  South  Dakota. 
(c)  Vacuum-filtration.— Best  examples:  Terry,  South  Dakota,  Lib- 
erty Bell  mine,  Colorado ;  Goldfield,  and  Virginia  City,  Nevada. 

Third  method. — For  ore  that  contains  both  coarse  and  fine  gold. 
The  coarse  gold  is  recovered  by  amalgamation,  and  a  large  part  or 
the  whole  of  the  tailing  from  the  plates,  is  re-ground  or  slimed  in 
pans  or  tube-mills.  The  ground  product  is  agitated  and  treated  by 


OF    THE    COMMON    METALS.  157 

decantation,  filter-pressing,  or  vacuum-filtration.     Concentration  is 
omitted.     Best  example :    El  Oro,  Mexico. 

Fourth  method. — For  ore  that  contains  fine  gold  only.  The 
crushed  ore  is  classified  into  sand  and  slime,  the  former  being 
leached.  Concentration  is  omitted.  A  feature  of  the  process  is  that 
the  ore  is  crushed  in  weak  cyanide  solution.  Best  example :  Mait- 
land  mill,  South  Dakota. 

Fifth  method. — For  ore  that  contains  gold  in  combination*  with 
tellurium.  The  ore  receives  an  oxidizing  roast  after  which  it  can  be 
treated  by  the  'first  method'.  As  the  result  of  roasting  some  of  the 
gold  may  take  the  form  of  shot,  beads,  or  coarse  particles  of  gold 
that  settle  and  must  be  caught  by  a  classifier,  riffle,  sluice,  concentrat- 
ing table,  or  amalgamated  plate.  If  ore  is  re-crushed  in  tube-mills, 
the  shot  or  coarse  gold  becomes  ground  so  fine  it  all  may  dissolve  in 
the  cyanide  solution.  Best  examples :  Cripple  Creek  and  Western 
Australia. 

40.     FIRST  METHOD  OF  CYANIDATION. 

This  was  the  first  method  put  in  practice  for  the  treatment  of  gold 
ores.  It  is  applicable  particularly  to  free-milling  gold  ores,  in  which 
the  gold  is  in  a  free  or  uncombined  state.  Also,  since  the  cyanide 
solution  acts  but  slowly  upon  gold,  it  is  necessary  when  using  this 
method,  that  the  gold  particles  be  extremely  fine;  otherwise  it  re-- 
quires  a  long  time  to  leach  the  ore.  When  coarse  gold  is  present  it 
it  first  caught  on  amalgamated  plates.  The  ore  is  crushed  preferably 
by  rolls  to  avoid  the  formation  of  slime,  which  would  interfere  with 
the  leaching.  The  crushed  ore  is  leached  in  vats  with  weak  cyanide 
solution,  and  the  gold  is  precipitated  from  the  filtered  solution  by 
passing  it  through  boxes,  treated,  and  obtained  in  the  form  of  gold 
ingots. 

Method  of  crushing. — A  comparison  between  stamps  and  rolls 
shows  the  former  to  be  durable,  less  expensive  to  install,  and  capable 
of  finely  crushing  ore  from  a  rock-breaker.  Where  stamps  are  used 
the  ore  is  wet  and  four  to  eight  tons  of  water  is  needed  per  ton  of  ore. 
They  make  a  greater  proportion  of  extremely  fine  material  or  slime 
in  crushing  than  rolls.  Where  ore  contains  talc  or  clay,  that  would 
cause  it  to  slime,  rolls  give  a  more  granular  product,  and  one  that  is 
more  easily  percolated  or  leached  by  the  cyanide  solution.  With 
good  rolls,  an  efficient  system  of  graded  crushing  can  be  used  and 
dry  crushing  is  preferred  for  the  following  reasons : 

(1)     The   ore   after  crushing  is  delivered  dry  to  the  leaching- 


158 


THE    METALLURGY 


tanks,  and  there  is  no  dilution  of  solution  as  when  treating  a  wet- 
crushed  ore  which  contains  10  to  20%  moisture. 

(2)  In  dry-crushing  water  is  not  needed.    This  is  an  important 
consideration  in  a  dry  country  where  water  is  scarce. 

(3)  A  uniform  bed  of  easily  percolated  ore  can  be  put  in  the 


OF    THE    COMMON    METALS. 


159 


leaching-tank,  the  ore  is  better  aerated,  and  the  oxygen  present  as- 
sists in  the  solution  of  the  gold,  as  shown  in  Eisner's  reaction. 

On  the  other  hand  the  cost  of  dry  crushing  beyond  40-mesh  is 
excessive. 

Leaching  tailing. — Fig.  65  and  66  represent,  in  sectional  elevation 


160 


THE    METALLURGY 


and  plan,  a  75-ton  plant  designed  to  treat  impounded  tailing. 
which  has  accumulated  from  a  mill  treating  gold  ore  of  a  kind 
suited  to  this  method.  The  tailing  contains  so  little  slime  that  only 
37%  passes  an  80-mesh  sieve;  and  the  gold  can  be  extracted  by 
four  days  leaching. 

Referring  to  plan.  Fig.  66,  &,  b,  &,  &,  are  the  leaching  vats  (some- 
times called  percolators)  the  most  important  units  of  the  plant,  and 
the  ones  to  which  the  others  are  accessory.  They  are  20  ft.  diam. 


Fig.    67.      WOODEN  LEACHING   VAT   SHOWING   FALSE   BOTTOM. 

by  6%  ft.  deep,  having  75  tons  capacity  above  the   filter-bottom. 
The  four  together  allow  four  days  treatment  for  each  charge. 

They  are  constructed  with  filter  bottoms,  as  shown  in  Fig.  67 
and  in  perspective  Fig.  68.  They  may  be  made  of  wood  or 
steel.  In  warm  countries,  like  Australia,  South  Africa,  or  Mexico, 
the  steel  vat  is  to  be  preferred ;  but  in  cold  countries,  where  it 
is  necessary  to  house  the  plant,  the  wooden  vat  gives  satisfaction. 
The  latter  is  cheaper  in  first  cost  and  easier  to  set  up,  though  the 
wood  absorbs  gold  solution.  The  steel  vat  on  the  other  hand  is 
less  liable  to  leak,  but  costs  more  to  maintain  in  requiring  a 


OF    THE    COMMON    METALS. 


161 


periodical  coat  of  protective  asphalt  paint  to  preserve  the  steel 
from  the  action  of  the  cyanide  solution.  A  wooJcn  tank,  a!so7  \\h_i 
requiring  it,  should  be  painted.  In  Western  America  wooden  vats 
predominate.  They  are  carefully  built  by  companies  who  make  this 
their  business.  Steel  vats  are  also  used.  Each  kind  has  advocates, 
but  in  general  the  steel  vat  is  preferred. 

Fig.  68  is  a  perspective  view  of  a  wooden  vat  with  the  filter- 
cloth  omitted.  This  shows  the  false-bottom  of  slats  and  the  hinged 
bottom-discharge  opening  through  which  the  exhausted  tailing  is 
shoveled  or  sluiced  out.  A  wooden  ring,  2  in.  high  and  21/->  thick, 


Fig.    68.      PERSPECTIVE  VIEW  OF  WOODEN  LEACHING  VAT. 

is  nailed  to  the  bottom  of  the  vat,  leaving  a  space  of  %  in.  between 
it  and  the  side.  Parallel  strips,  one  inch  high,  are  nailed  a  foot 
apart  upon  the  bottom,  and  across  these  1  by  4  in.  strips  or  slats 
are  laid  with  1-in.  space  between.  Upon  this  false-bottom  cocoa- 
matting  is  spread,  and  over  it  8-oz.  canvas  filter-cloth  cut  12  in. 
larger  in  diameter  than  the  vat.  The  edges  of  the  cloth  are  held 
down  by  a  rope  laid  upon  the  canvas  and  driven  into  the  %-in.  space 
between  the  staves  and  the  wooden  ring.  This  way  of  securing 
the  canvas  is  shown  in  further  detail  in  Fig.  71. 

Fig.  69  gives,  in  plan  and  section,  the  construction  of  another 
form  of  filter-bottom  for  a  wooden  leaching-vat.  Wooden  rings  of 
graded  heights  are  nailed  to  the  bottom  of  the  vat  and  upon  these 
are  nailed  radial  pieces  bored  with  %-in.  holes.  This  arrangement 
of  a  sloping  bottom  facilitates  the  discharge  of  tailing  when  sluiced 


162  THE    METALLURGY 

or  hosed  out  of  the  vat.  The  wooden  ring,  in  Fig.  67  is  omitted  and 
instead,  the  joint  is  made  by  nailing  a  wooden  strip  around  the 
tank  so  as  to  hold  the  canvas  tight  against  the  staves. 

Fig.  70  and  71  represent,  in  plan  and  in  elevation  respectively, 
the  construction  of  a  steel  vat  having  a  perforated  board  bottom. 


Fig.    69.       PLAN  AND   SECTION  OF  WOODEN  LEACHING  VAT. 

A  ring  of  flat  iron  %  by  2l/2  in.  is  rivetted  to  the  side  of  the  tank, 
with  space-thimbles  to  hold  it  %  in.  from  the  side.  The  cleats  that 
sustain  the  false-bottom  are  2  in.  high  by  I1/)  in.  wide.  The  1-in. 
bottom-boards  are  bored  with  %-in.  holes  and  screwed  to  the  cleats. 
As  in  the  case  of  the  wooden  vats,  the  thick  stiff  cocoa-matting  is 
laid  upon  the  false-bottom,  and  covered  with  a  filter-cloth  of  8  to 


OF    THE    COMMON    METALS. 


163 


10-oz.   canvas.     The  edges  are  calked  with   a  %-in.  rope  into  the 
%-in.  space,  as  shown  at  g  in  the  sectional  view,  Fig.  71. 


(b) 


Fig.    70.      PLAN  OF  STEEL  LEACHING  VAT. 

Leaching  vats  vary  in  size  from  16  to  50  ft.  in  diam.  and  4  to  9 
ft.  in  depth.  The  shallow  ones  are  for  the  more  finely-ground  sand. 
The  tank  is  generally  made  to  hold  one  day's  supply  of  ore  to  insure 


Fig.   71.      SECTION  OF  STEEL  LEACHING  VAT. 

uniform  work  in  the  mill.    Hence  the  number  of  tanks  indicates  the 
number  of  days  of  treatment. 

The  exhausted  sand  of  the  vat  may  be  shoveled  out  through  side 
or  bottom  openings,  or,  where  water  is  abundant,  may  be  washed  by 


164  THE    METALLURGY 

a  hose  into  a  launder  as  shown  in  s  Fig.  65,  and  thus  conveyed  to  the 
dump. 

Fig.  72  represents  a  side-discharge  door  which  is  bolted  to  the 
outside  of  the  vat,  and  which  can  be  quickly  opened  or  closed.  The 
joint  between  the  door  and  frame  is  made  tight  with  a  rubber  gasket. 
Fig.  73  is  a  bottom-discharge  valve.  Opening  upward,  it  is 
conveniently  operated  from  the  platform  above  the  tank  when  the 
charge  is  to  be  sluiced  out.  It  is  securely  bolted  to  the  bottom, 
and  is  self-sustaining.  For  a  large  tank  four  such  valves  may  be 
used.  The  opening  in  each  is  10  in.  diam.  In  Fig.  68  is  shown  a 
drop-bottom  door  that  opens  from  below. 

At  Z  Fig.  65,  is  the  charging  platform  or  bridge  (not  shown  in 
Fig.  66),  grated  where  it  is  over  the  vats,  with  2-in.  openings.  The 


Fig.   72.      SIDE   DISCHARGE   DOOR. 

tailing  is  brought  by  a  wagon  having  movable  bottom-planks,  and 
dumped  upon  the  grating  from  which  it  falls  lightly  into  the  vat. 
When  full  the  surface  is  leveled  with  a  hoe.  At  y  is  a  platform  set 
41/2  ft.  below  the  top  of  the  vat  for  convenience  in  leveling  the  ore 
and  in  adjusting  the  supply-valves.  At  a  and  a'  Fig.  66  are  the 
strong  and  weak-solution  stock-tanks,  one  of  which  contains  5  Ib. 
potassium  cyanide  solution,  the  other  a  weak  solution,  0.1%,  or  2  Ib. 
per  ton.  These  tanks  deliver  solution  to  the  leaching-vats  by  pipe 
c  for  the  strong  and  by  c/  for  the  weak  solution.  The  filtrate 
from  the  vats  goes  by  a  double  launder,  d  for  the  strong  and  d'  for 
the  weak  solution.  These  deliver  by  the  cross-launder  j  and  /'  to 
the  strong  gold-solution  tank  h  and  the  weak  gold-solution  tank  li' '. 
Charging  the  vat. — Where  tailing  has  been  accumulated  in 
dumps,  it  is  coarser  at  the  point  of  discharge  from  the  mill  and 


OF    THE    COMMON    METALS. 


165 


finer  toward  the  further  limit,  and  to  treat  the  accumulation,  the 
two  must  be  mixed.  They  may  be  mixed  in  the  reservoir  by  plowing 
and  breaking  up  the  lumps  with  a  disk-harrow.  When  the  material 
has  become  pulverized  and  dry  it  is  scraped  into  heaps.  The  coarse 
and  fine  are  together  shoveled  into  a  wagon,  and  the  mixing  is 
completed  when  the  tailing  is  dropped  from  the  wagon,  through  the 
grating,  into  the  leaching  vat.  The  material  should  be  loaded  evenly 
into  the  vat  and  should  fill  it,  so  that  when  settled  by  the  wash-water, 


Fig.   73.      BOTTOM  DISCHARGE   VALVE. 

the  level  will  be  10  in.  below  the  top.  This  gives  room  for  the 
wash-water  and  cyanide  solutions.  Since  the  tailing  is  acid  and  the 
acid  would  destroy  potassium  cyanide,  it  is  necessary  to  add  a 
certain  proportion  of  slaked  quick-lime  when  charging  into  the  vat. 
Leaching  the  ore. — A  drop-pipe  from  the  strong-solution  pipe 
c  extends  just  below  the  filter  bottom;  and  by  this  a  0.25%  KCN 
solution  is  introduced.  It  rises  through  the  ore  and  is  shut  off 
when  the  ore  is  covered  2  or  3  in.  deep.  The  charge  is  allowed  to 
remain  in  this  condition  a  fixed  time  after  which  the  discharge- 
valve  at  the  bottom  of  the  tank  is  opened  and  the  solution,  at  first 
weakened  by  the  water  it  has  taken  from  the  charge,  is  run  to  the 


166 


THE    METALLURGY 


weak-solution  gold-tank.  The  strength  steadily  increases  and  when 
above  0.1%  (the  grade  of  weak  solution)  the  solution  is  turned 
into  the  strong-solution  gold-tank. 

Percolation  now  proceeds,  solution  being  added  above  the  surface 
of  the  charge  in  a  succession  of  washes  between  each  of  which  the 


OF    THE    COMMON    METALS. 


167 


solution  is  allowed  to  sink  beneath  the  surface,  to  draw  air  into 
the  charge.  Weak  solution  is  now  admitted  above  the  charge  to 
displace  strong,  and  as  the  effluent  weakens  it  is  turned  into  the 
weak-solution  tank  taking  care  to  prevent  a  needless  accumulation 
of  Aveak-solution,  by  using  as  little  as  possible. 

Displacement  of  one  solution  of  the  charge  by  another  does 
not  take  place  uniformly ;  the  compactness  of  the  charge  is  irregular 
and  the  more  permeable  parts  allow  the  quicker  advance  of  the 
solution  last  added,  causing  it  to  mix  more  or  less  with  that  preceed- 
ing.  It  has  been  found,  in  fact,  that  of  weak  solution  l1/^  to  2 
times  the  theoretical  quantity  is  needed  to  displace  the  strong. 
Different  charges  percolate  at  different  rates,  and  no  hard-and-fast 
rule  can  be  established  for  handling  tailing.  Each  charge  should 


Fig.    In.      PERSPECTIVE    VIEW   OF   ZINC    BOXES. 

be  studied  by  itself,  and  frequent  samples  of  the  out-going  solution 
must  be  chiefly  relied  on  to  make  sure  of  proper  extraction. 
The  cycle  of  treatment  of  a  75-ton  charge  is  as  follows : 

Hours. 

Charging,  or  filling  the  vat  with  tailing 9 

Leveling    charge    and    saturating    with    strong 

solution    3 

Percolation  until  effluent  becomes  0.18%  KCN.   20 
Continued  percolation  with  strong  solution ...   20 

Displacing  strong  solution  with  weak 29 

Displacing  weak  solution  with  water 13 

Sluicing  out  the  exhausted  tailing 3 


Total  time  (4  days  1  hour) 97 

The   solution   from  the   gold-solution   tanks   enters   the   zinc   or 


168 


THE    METALLURGY 


extractor-boxes  of  which  three  receive  the  strong  and  two  the  weak 
solution.  Fig.  74  gives  a  plan  and  elevation  of  a  zinc-box,  and 
Fig.  75  a  perspective  view. 

As  shown,  each  set  of  boxes  contains  seven  compartments,  each 
12  by  15  by  24  inches  in  size.  The  compartments  have  perforated 
false-bottoms  of  sheet-iron  or  wire-cloth  that  sustain  the  zinc-shaving 
with  which  they  are  filled.  The  partitions  are  set  alternately  up 
and  down,  to  compel  an  upward  flow  of  the  gold-bearing  solution 
through  the  shaving,  and  to  bring  it  intimately  in  contact  with 
the  zinc  to  insure  the  precipitation  of  the  gold  upon  the  surface 
of  the  zinc.  At  a  in  the  side  elevation  (Fig.  74)  the  solution  enters 
the  box  through  a  pipe  indicated  at  the  left,  passes  through  all 
the  compartments,  flows  over  the  last  partition,  and  discharges 
through  a  down-turned  pipe  into  the  sump-tank.  The  boxes  are 
set  at  a  grade  of  %  in.  to  the  foot. 

Precipitation. — The  strong  and  weak  solutions,  leaving  the 
leaching  vats,  are  gathered  in  the  respective  gold-tanks.  These 


Fig.    76.      FLOATING  HOSE  FOR  GOLD   SOLUTION  TANK. 

tanks  accumulate  the  gold-bearing  solutions  from  the  leaching-vats, 
and  provide  a  uniform  discharge  into  the  zinc-boxes.  Since  the 
solutions  contain  a  little  flocculent  precipitate,  which  if  allowed 
to  enter  the  zinc-boxes  would  seriously  interfere  with  the  precipita- 
tion of  the  gold  on  the  zinc,  they  are  allowed  to  settle  while  the 
clear  supernatant  solution  is  drawn  from  the  tank  by  a  2-in.  floating- 
hose,  as  shown  in  Fig.  76.  This  hose  is  attached,  at  its  free  end,  to 
a  float  /,  which  may  be  a  5-gal.  oil-can  painted  with  asphalt  paint. 
The  high  end  of  the  hose  is  thus  sustained  slightly  below  the  surface. 
The  solution  enters  behind  a  partition  &,  so  that  it  deposits  its 
sediment  at  the  bottom  of  the  tank.  The  accumulated  sediment  is 
occasionally  drained  from  the  tank  through  the  plug-opening  s 
close  to  the  bottom.  Solutions  pass  through  the  zinc-boxes  in  a 
regulated  flow  into  the  sumps,  the  gold  being  precipitated  on  the 
zinc-shaving,  which  permits  the  free  passage  of  the  solution. 

Zinc  shaving  is  either  bought  prepared,  or  preferably  is  made  at 


OF    THE    COMMON    METALS.  169 

the  works.  When  freshly  made,  it  is  more  efficient.  To  make  it,  a 
sheet  of  zinc  is  wound  around  a  mandrel  in  a  lathe,  and  the  edge  of 
the  sheet  is  soldered  down.  A  side-cutting  tool  is  used  to  cut  the 
shavings  which  are  V120o  in-  thick,  1/32  in.  wide,  and  several  feet  long. 
All  the  compartments  of  the  zinc-box,  except  the  last,  are 
compactly  but  uniformly  filled  with  the  shaving,  the  last  compart- 
ment serving  as  a  settling  box.  Strong  solution  flows  through  the 
three  strong-solution  zinc-boxes  at  the  rate  of  one  ton  per  hour. 


Fig.      77.      ZINC  LATHE. 

The  gold  precipitates  chiefly  in  the  compartments  first  traversed,  so 
that,  by  the  time  the  solution  reaches  the  seventh  one,  no  discolor- 
ation of  the  shaving  is  to  be  noticed,  and  the  barren-solution  can 
go  to  the  sump-tank  /.  The  deposit  on  the  zinc  has  a  brownish, 
then  a  grayish-black  hue.  As  it  increases  in  quantity,  the  shaving  in 
the  first  compartment  becomes  soft  and  stringy,  and  both  precipitate 
and  small  pieces  of  the  zinc  settle  through  the  screen  to  the  bottom 
of  the  compartment.  As  the  shaving  in  the  first  compartment 
settles  down  it  is  usual  to  replenish  it  with  zinc  from  the  last 
compartments,  which  in  turn  are  filled  with  fresh  shaving. 

Before  putting  the  shaving  into  the  boxes  it  is  usual  to  dip  it 


170  THE    METALLURGY 

in  a  solution  of  lead  acetate.  Lead  deposits  as  a  film  on  the  shaving, 
forming  a  zinc-lead  couple  that  is  more  active  than  the  zinc.  Where 
copper  is  present  in  the  solution  selective  precipitation  is  practised. 
This  consists  in  running  into  it,  just  before  it  enters  the  zinc-boxes, 
fresh  cyanide  solution  to  raise  its  grade.  In  this  way  copper  is  held 
in  solution,  less  of  it  precipitating  upon  the  shaving.  From  95  to 
99%  of  the  gold  precipitates  on  the  zinc.  Since  the  barren  solution 
is  pumped  back  to  the  stock-tanks,  the  residual  gold  is  not  lost. 

The  weak  solution,  passing  through  the  two  boxes  k'  A:'  Fig.  66 
in  series,  deposits  the  gold  with  more  difficulty  and  less  completely 
than  the  strong.  The  solution  goes  through  the  first  box,  then  the 
second,  or  through  14  compartments  in  all,  to  precipitate  as  much  of 
the  gold  as  possible.  It  then  flows  into  the  tank  I' . 

The  clean-up. — This  is  made  monthly  or  bi-monthly,  according 
to  the  bulk  of  the  precipitate  to  be  treated,  and  the  need  of  realizing 
values  for  operating  expenses. 

At  the  time  of  the  clean-up,  only  one  zinc-box  is  taken  care  of  at 
a  time,  the  flow  continuing  in  the  other  boxes.  The  flow  of  gold 
solution  to  this  box  is  stopped,  and  water  is  run  in  to  displace 
the  solution  contained.  Beginning  in  the  first  compartment,  the 
shaving  is  agitated  in  the  water  for  five  minutes  with  the  hands 
protected  by  rubber  gloves.  This  is  not  done  roughly,  for  the  brittle 
shaving  would  be  unnecessarily  broken  and  the  water  would  be 
black  with  the  floating  precipitate.  The  plug  in  the  side  (see  k  in 
the  cross-section  Fig.  74)  is  gradually  withdrawn,  and  the 
accumulated  slime  and  water  allowed  to  flow  into  the  launder  //. 
The  double  line  at  the  side  of  the  zinc  boxes,  in  Fig.  66,  shows  this 
launder.  It  connects  by  a  cross-launder  n  and  a  main  one  m  to  the 
acid  tank  o,  6  ft.  diam.  The  plug  is  replaced  and  the  compartment 
is  again  filled  with  water.  The  zinc  again  is  rinsed  and  rubbed,  and 
the  loosened  precipitate  once  more  drawn  off.  About  three  such 
washes  free  the  shaving  from  precipitate  and  short-zinc.  The 
compartments  are  thus  successively  cleaned  up,  and  the  shaving 
from  each  compartment  successively  is  moved  toward  the  head, 
and  in  the  last  compartment,  where  needed,  replaced  by  fresh 
shaving.  Finally,  the  launder  is  cleaned  with  a  hose,  and  everything 
washed  into  the  acid-tank  o. 

When  all  the  boxes  have  been  cleaned,  the  precipitate  is  allowed 
to  settle  a  short  time  in  the  acid-tank  and  the  supernatant  liquid 
is  syphoned  into  the  8-ft.  settling  tank  p.  In  this  larger  tank  the 
particles  of  precipitate  have  opportunity  to  settle,  and  to 
be  recovered  subsequently  in  the  filter-press.  The  acid-tank  is  stirred 


172  THE    METALLURGY 

by  hand  with  a  wooden  hoe,  or  preferably  a  power-driven  agitator 
in  constant  motion.  This  insures  a  thorough  agitation  of  the  sludge 
and  precipitated  in  the  acid-treatment  now  to  be  described. 

The  acid  treatment. — Upon  the  watery  slime  about  30  Ib.  of 
sulphuric  acid  is  poured.  This  acts  upon  the  short-zinc  and  produces 
a  violent  effervescence.  After  subsidence  the  whole  is  stirred. 
When  again  the  action  abates,  15  Ib.  hot  water  and  the  same  amount 
of  acid  is  added  with  occasional  stirring.  This  is  repeated  until 
further  addition  of  acid  produces  little  effervescence.  Then  the 
mixture  is  allowed  to  stand  two  hours,  and  a  portion  is  tested  with 
more  acid  to  see  that  decomposition  is  complete.  The  total  time 
for  the  operation  is  four  to  six  hours. 

Filter-pressing. — The  black  mixture,  containing  zinc  sulphate 
in  solution,  is  diluted  with  hot  water  to  within  a  few  inches  of  the 
top  of  the  tank.  The  whole  content  is  stirred  and  then  pumped 
through  the  filter-press  r  Fig.  66.  The  tank  is  washed  and  the 
washings  also  are  pumped  through  the  press.  Finally  the  residue 
in  the  press  is  washed  with  hot  water  to  entirely  remove  the  zinc 
sulphate. 

Fig.  78  represents  the  filter-press  used.  It  consists  of  a  series 
of  flat  cast-iron  frames  18  in.  square.  The  frames  are  recessed, 
and  between  them  are  held  canvas  filter-cloths  when  in  position 
for  filtering.  At  the  left-hand  corner  of  the  illustration,  a  dotted 
line  on  the  cross-section  of  the  frame  indicates  how  the  filter-cloth 
rests  against  the  filter-frame.  The  filter-cloths,  for  use,  are  covered 
with  filter-paper,  so  that  the  precipitate  does  not  touch  the  cloth. 
This  paper  is  burned  after  use,  and  the  ash  is  mixed  with  the 
precipitate.  The  slime  from  the  pulp  enters  the  press  under  pressure 
through  a  pipe  and  valve,  as  shown  at  the  left.  The  filtrate 
escapes  by  a  row  of  bronze  cocks,  one  on  each  frame,  into  a 
launder,  and  thence  flows  to  the  settling  tank  p  Fig.  66.  Leaning 
against  the  launder  are  to  be  seen  two  of  the  filter-frames.  These 
show  the  grooved  surfaces  along  which  the  liquid,  after  passing 
through  the  filter-cloth,  reaches  the  outlet  hole  drilled  parallel  with 
the  frame  and  leading  to  the  bronze  cocks.  In  case  a  filter-cloth 
breaks,  so  that  pulp  begins  to  issue  from  any  cock,  this  can  be 
shut  off.  The  mixture  enters  through  the  centers  of  the  plates 
and  distributes  itself  sideways  into  the  recesses  between  the  filter- 
cloths. 

The  entire  precipitate,  having  been  transferred  to  the  press, 
while  in  this  position  is  washed  with  water  under  pressure.  The 
water  is  followed  by  compressed  air  to  dry  the  precipitate,  which, 


OF    THE    COMMON    METALS.  173 

after  this,  is  ready  to  discharge.  To  discharge  the  press, 
the  tightening-screw  at  the  right  is  slackened ;  the  follower  is  drawn 
back,  the  frames  are  successively  separated.  The  grayish-black 
residue  in  the  recesses  of  the  frame  containing  20%  or  less  of 
water,  drops  into  a  drying-pan  placed  beneath  the  press  to  receive  it. 
Drying  and  final  treatment  of  the  precipitate. — The  product, 
still  damp,  is  transferred  to  pans  44  by  24  in.  by  4  in.  deep.  A 
pan  is  slid  into  a  cast-iron  muffle  and  heated  until  the  precipitate 
is  dry  and  then  heated  to  an  incipient  red.  It  is  then  removed, 
allowed  to  cool,  and  the  weighed  contents  cautiously  mixed  with 
half  its  weight  of  a  flux  composed  of  4  parts  borax,  2  parts  soda, 
and  1  part  sand.  At  the  Standard  plant,  Bodie,  where  these 


Fig.   79.     CENTRIFUGAL  PUMP. 

proportions  are  used,  the  precipitate  contains  6%  silver  and  9.5% 
gold.  A  clean  and  fusible  slag  is  obtained  by  the  use  of  the  flux. 
Since  the  product  is  light  and  dusty,  care  must  be  taken  in  handling 
it,  and  for  fusing,  it  must  be  put  cautiously  into  the  No.  60  plumbago 
(3-gal.)  crucible.  The  melting  is  performed  in  the  furnace  shown 
at  the  right  in  Fig.  106. 

The  crucibles  are  placed  in  a  wind-furnace  and  packed  in  coke. 
The  melting  is  done  as  described  under  silver  milling.  The  molten 
metal  is  stirred  in  the  crucible,  then  poured  into  conical  molds ; 
and  on  cooling,  the  slag  is  removed  and  the  gold  remelted  into 
an  ingot. 

At  v  Fig  66  is  a  vacuum  pump  by  which  air  and  solution  are 
withdrawn  from  beneath  the  filter-bottoms  of  the  leaching-tanks 


174  THE    METALLURGY 

through  the  pipes,  e,  e,  and  discharged  by  the  pipe  u  to  the  stock- 
tanks  a  and  a'.  The  solution  that  has  lost  strength  by  mixing  with 
weak  solution  or  water  is  increased  to  standard  strength  in  the 
stock-tank. 

The  position  of  the  engine  which  drives  the  overhead  line- 
shaft  is  indicated  by  g  in  Fig  66,  while  w  Fig.  65  is  the  position 
of  the  centrifugal  pump  driven  from  the  same  shaft  by  which  solution 
from  the  sumps  /  and  I'  is  returned  to  the  stock-tanks  a  and  a'. 
Fig.  79  is  a  view  of  such  a  pump.  The  solution  enters  through 
the  elbow  at  the  left  and  delivers  from  the  discharge  at  the  top. 
The  pump  has  no  inside  valves  to  get  out  of  order,  and  can  be 
run  at  a  velocity  of  500  or  more  revolutions  per  minute. 

A  complete  plant. — Fig.  80  is  an  elevation  and  Fig.  81  a  view 
of  a  plant  for  the  treatment  of  ore  by  the  first  method  directly 
from  the  mine.  This  has  been  designed  by  the  Allis-Chalmers  Co., 
Milwaukee,  Wisconsin. 

As  shown  in  Fig.  81,  the  buildings  include  a  storage-building  for 
the  ore-bin,  a  coarse-crushing  and  drying-room,  and  a  fine-crushing 
building,  all  of  which  are  for  the  crushing  plant  and  the  preparation 
of  the  ore.  The  building  used  for  the  cyanide  plant  is  the  largest 
of  all,  and  adjoining  it,  shown  in  the  foreground,  is  the  power 
house.  At  the  right,  and  in  front  of  the  cyanide  plant,  the 
tailing-dump  is  indicated. 

The  ore  from  the  mine  enters  at  a  high  level  and  is  discharged 
into  the  ore-bin  a  Fig.  80.  From  the  bin  it  is  withdrawn  through 
a  gate  to  a  pile  on  the  floor  near  the  mouth  of  the  Gates  crusher, 
size  2.  The  entire  ore,  including  fine  and  coarse,  is  regularly 
shoveled  into  the  crusher,  and  broken  to  1-in.  size,  the  operation 
being  confined  to  the  10-hour  day-shift.  Since  we  desire  to  crush 
ore  into  a  product  as  granular  as  possible  that  it  all  may  be 
leached,  dry  crushing  is  here  practised.  The  finer  crushing  is 
preceded  by  drying  the  ore  in  order  that  it  may  be  screened  during 
the  crushing.  It  is  accordingly  passed  through  a  cylindrical  dryer 
c  18  ft.  long  with  ends  36  and  44  in.  diam.  respectively.  From 
the  dryer,  the  ore  goes  to  a  special  fine  crusher  H,  where  it  is 
reduced  to  y±-in.  size.  This  discharges  into  an  elevator  d,  which 
carries  it  to  an  inclined  double  shaking-screen,  so  arranged  that 
the  ore-stream  may  be  diverted  to  either  screen,  thus  guarding 
against  loss  of  time  in  case  of  accident  or  repairs.  This  screen  is 
simple  in  action,  and  low  in  first  cost,  and  it  easily  treats  the 
desired  quantity  (5  tons  per  hour)  screening  it  to  8-mesh  size. 
The  over-size  of  the  screen  slides  back  to  the  special  fine  crusher, 


176  THE    METALLURGY 

while  the  under-size  goes  to  crushing  rolls  set  for  fine  crushing. 
The  product  from  the  rolls  is  raised  by  the  elevator  /  to  another 
double  30-mesh  shaking-screen,  the  reject  from  which  passes  back 
to  the  rolls,  while  the  under-size  goes  to  a  storage  bin  k.  It  will 
be  noticed  that  in  this  system  there  is  no  automatic  feeding  of 
ore,  and  no  storage  after  coarse-crushing.  Simplicity  is  obtained 
by  the  use  of  a  fine  crusher  in  place  of  rolls,  and  by  the  use  of 
the  inclined  screen.  On  the  other  hand,  the  system  requires  the 
more  costly  side-hill  site,  and  at  the  same  time,  does  not  escape 
the  need  of  elevators. 

The  product  when  crushed  is  drawn  from  the  bin  into  scoop- 
cars  standing  on  platform  scales;  and  the  content  of  each  car 
is  weighed.  A  suspended  tract  /,  with  turn-tables  and  tracks  at 
right  angles,  is  constructed  over  the  leaching  or  percolating-vats, 
I,  ?,  by  means  of  which  the  cars  may  be  dumped  directly  into  the 


Fig.    81.      VIEW   OF   FIFTY-TON   CYANIDE   PLANT. 

vats,  the  ore  being  leveled  with  a  hoe.  There  are  six  leaching-vats, 
each  of  a  capacity  of  50  tons,  for  a  six-day-cycle  treatment  each 
vat  being  24  ft.  diam.  by  3  ft.  deep. 

Instead  of  dumping  the  ore  directly  into  the  vat  from  a  height, 
which  would  tend  to  pack  it  in  spots,  the  better  way  proposed  is 
to  have  a  hopper  centrally  placed  over  each  tank  just  below  the 


OF    THE    COMMON    METALS. 


177 


track-level.  An  adjustable  spout  leads  from  this.  By  it  the  ore 
from  the  hopper  may  be  more  gently  directed  to  any  part  of  the 
vat,  and  kept  in  a  looser,  open,  and  uniform  condition  for  leaching. 
The  strong  and  weak-solution  stock-tanks  are  situated  at  the 
top  of  the  building,  and  supply  solution  to  the  leaching  vats  for 
ore  treatment.  There  are  two  vacuum-tanks  used  in  connection 
with  the  vacuum-pump.  Upon  connecting  them  to  the  leaching-vats 
by  a  pipe  that  enters  below  the  false-bottom,  the  operation  of 
percolation  is  hastened,  especially  toward  the  last,  when  the  ore 
becomes  settled  and  compacted  after  the  washes.  When  the  vacuum- 
tank  is  filled,  the  solution  is  stopped  and  the  tank  is  discharged  into 


Fig.    82.      THREE-COMPARTMENT   SPITZKASTEN. 

the  gold-solution  tanks  m  or  m'.  From  these  the  solution  flows 
through  the  zinc-boxes  to  the  respective  strong  and  weak-sumps 
//  and  n".  From  the  sumps  the  solutions  are  pumped  back  to  the 
storage  or  stock-tanks  above. 

The  tailing,  still  retaining  some  16  to  18%  water,  is  shoveled 
through  bottom-discharge  doors  into  tram-cars  and  run  to  the 
tailing  dump.  The  treatment  of  the  zinc-box  precipite  has  been 
described. 

41.     CLASSIFIERS   AND    SETTLING   TANKS. 

Sand  and  slime,  mixed  and  in  suspension  in  water  as  in  ore 
pulp,  may  be  separated  under  free-settling  conditions  by  devices 
called  classifiers.  The  pulp  passes  through  the  classifiers,  from 
the  entrance  to  the  discharge  in  a  current,  and  the  heavier  and 


178  THE    METALLURGY 

larger  particles  (the  sand)  separate  by  gravity,  while  the  finer 
portion  (the  slime),  still  in  suspension,  passes  out  through  the  over- 
flow with  the  water.  The  products  are  therefore  termed  the  'over- 
flow' and  the  'under-flow, '  or  spigot-discharge.  Some  of  these 
classifiers  have  a  supply  of  water,  'hydraulic  water',  rising  from 
below,  just  sufficient  to  keep  the  slime  in  suspension  until  it  escapes 
in  the  overflow,  while  the  sand  settles  against  the  rising  current 
and  escapes  in  the  spigot-discharge. 

Of  the  classifiers  that  do  not  require  hydraulic  water  we  here 
enumerate  two  kinds:  spitzkasten,  and  classifying  cones. 

Fig.    82    shows,    in    plan    and    sectional    elevation,     a    three- 


FIG.  83.       CALLOW   CLASSIFYING   CONE. 

compartment  spitzkasten.  The  current  enters  by  a  launder  at  a, 
and  traverses  the  three  compartments  successively,  the  overflow 
escaping  by  the  launder  I.  The  current  widens,  and  as  it  enters 
each  larger  compartment,  flows  more  slowly,  and  drops  successively 
the  finer  sand.  The  sand,  as  it  settles,  escapes  by  the  bent  pipes  s 
called  goosenecks.  Raising  thus  the  end  of  the  pipe  diminishes  the 
velocity  of  the  outflow,  and  less  water  escapes  with  the  sand.  We 
get  three  sizes  or  grades  of  sand  from  the  apparatus  shown :  the 
coarse  from  compartment  1,  medium  from  compartment  2,  and 
the  fine  from  compartment  3.  In  cyaniding  where  it  is  sought  to 


OF    THE    COMMON    METALS. 


179 


obtain    only   two   products,    a    single    compartment    in    a    classifier 
suffices. 

Fig.  83  is  a  perspective  view  of  a  Callow  classifying-cone, 
operating  on  the  principle  of  the  spitzkasten.  The  flowing  pulp 
enters  by  launder  to  the  center  of  the  cone  where  it  is  received  into 
a  short,  vertical,  12-in.  cylinder.  This  deflects  the  flow  vertically 
downward.  Rising  again,  it  overflows  evenly  around  the  rim,  and 
being  caught  in  the  circular  launder  at  the  rim,  flows  away  by 
the  spout  shown  at  the  front.  The  gooseneck  coming  from  the 


Fig.    84.      DOUBLE-CONE   CLASSIFIER. 

apex  of  the  cone  is  extended  into  a  rubber  hose.  The  end  can 
be  adjusted  to  alter  the  pressure  of  the  outflowing  sand-pulp. 
The  pipe  is  provided  with  a  side  valve,  through  which  water  can  be 
forced  and  an  outlet  at  the  apex,  to  clear  it  when  choked. 

Of  the  classifiers  using  hydraulic  water,  Fig.  84  is  an  example. 
It  is  a  double  sheet-metal  cone  with  a  cast-iron  sorting-chamber 
K  against  an  upward  current  of  water  admitted  at  E,  and  escaping 
by  the  spigot  at  H,  The  slime  and  fine  sand,  flowing  downward  to 
M  are  deflected  by  the  adjustable  cone  L,  caught  by  the  rising 
water-current,  and  carried  upward  between  the  cones,  escaping 
over  the  edge  of  the  outer  cone  at  the  level  C.  The  overflow  is 


180  THE    METALLURGY 

caught  in  the  circular  launder  and  escapes  by  the  spout  D.     The 
cast-iron  cone  L  is  adjusted  by  the  hand- wheel  0. 

Sometimes  the  inner  cone  is  omitted.  The  apparatus  then 
becomes  a  settling-cone  with  a  rising  current  of  water.  These  are 
not  really  satisfactory  in  their  action.  Hydraulic  water,  rising 
through  a  sorting-column  that  enlarges  rapidly  in  sectional  area 
decreases  in  velocity.  If  the  current  enters  the  classifier  in  a 
horizontal  stream,  the  heavier  grains  settle  out  and  light  ones 
overflow.  Heavy  grains  escape  in  the  sorting  column,  on  the  steep 
sides  of  the  cone  and  from  a  bank  there.  Occasionally  the  bank 
falls  in  a  mass  into  the  spigot-discharge  where  it  does  not  belong, 
or  remains  and  hinders  the  action  of  the  apparatus.  To  obviate 
this  defect  the  sides  of  the  cone  should  incline  55°  at  least.  If 
a  pointed  box  is  used  instead  of  a  cone,  the  bank  is  apt  to  form  at 
the  corners.  The  cone,  being  made  of  metal,  has  smooth  sides 
and  no  corners  for  a  bank  to  begin  to  form.  In  a  plain  cone,  using 
hydraulic  water,  it  is  better  to  use  a  pulsating  current.  This 
descends  to  the  bottom,  breaks  up  the  bank,  and  agitates  the  pulp, 
but  the  expedient  has  the  drawback  that  slime  enters  the  spigot- 
discharge  with  the  sand. 

42.     SECOND  METHOD  OF  CYANIDATION. 

Where  the  ore  carries  gold  and  pyrite,  a  stamp-mill,  such  as  shown 
in  Fig.  44,  is  employed.  The  coarse  gold  is  recovered  by  passing 
the  crushed  pulp  from  the  battery  over  amalgamated  plates,  and 
the  heavy  sulphide  (principally  pyrite)  is  removed  by  concentration 
on  Frue  vanners  or  other  concentrating  tables.  The  tailing  from 
the  mill  is  classified  into  sand  and  slime.  The  sand  is  treated  by 
leaching  as  in  the  first  method,  and  the  slime  by  one  of  three 
methods:  (a)  Decantation.  (b)  Filter-pressing,  (c)  Suction- 
filtration. 

The  plant,  for  the  cyanidation  of  tailing  is  called  the  cyanide 
plant;  and  this  may  be  divided  into  a  sand-plant  and  a  slime  plant. 
When  a  stamp-mill  is  provided  with  an  added  building  for  treating 
tailing  by  cyaniding,  such  an  addition  is  called  a  cyanide-annex. 

Plant  using  decantation. — The  practice  in  South  Africa  today 
has  been  brought  to  a  high  state  of  development.  At  first,  ores 
containing  $9  gold  per  ton  were  milled  by  wet-stamping  and 
amalgamating.  The  free  coarse  gold  was  recovered,  but  the  tailing, 
still  containing  $3.50  per  ton,  was  accumulated  in  extensive  deposits 
and  retained  behind  dams.  Cyanidation  was  first  applied  on  the 
accumulated  tailing  which  was  shoveled  into  cars  and  hauled  up 


OF    THE    COMMON    METALS. 


181 


inclined  tramways  to  large  leaching  tanks.  This  practice,  so  long 
as  the  impounded  tailing  lasted,  was  comparatively  simple ;  but, 
when  the  reserve  was  exhausted  it  was  necessary  to  devise  a  method 
for  treating  tailing  as  it  came  from  the  mill. 

Fig.  85  is  a  diagram  illustrating  how  this  is  done.  The  tailing 
from  the  stamps,  having  a  value  of  $9.60  per  ton,  is  neutralized 
with  milk  of  lime  at  the  launder  a.  and  enters  the  receiving  tank 
/;,  where  it  is  kept  agitated  by  a  4-arm  stirrer  mounted  on  a 
vertical  shaft.  A  centrifugal  pump  forces  it  by  the  pipe  c,  to  the 
pointed-box  classifier  d,  which  takes  out  a  mixture  containing  about 
10%  coarse  sand  and  $25  per  ton  concentrate,  this  being  delivered 
to  the  tank  g  by  a  distributor  /.  This  is  shown  more  clearly  in  the 


Fig.  85.   CYANIDE  PLANT  FOR  DOUBLE  TREATMENT. 


illustration  of  another  plant,  Fig.  86,  where,  in  the  foreground,  is 
a  settling  tank  arranged  with  a  Butters  distributor.  This  is  a 
horizontally  revolving  device  supported  by  a  vertical  shaft  on 
which  the  distributor  revolves.  The  pulp  from  the  classifier  is 
carried  in  a  launder,  shown  at  the  right,  to  the  central  hopper  of 
the  distributor  from  which  radial  pipes  discharge  the  pulp  into 
the  vat.  Each  pipe  is  bent  at  the  end  so  that  the  reaction  of  the 
escaping  pulp  sets  the  distributor  in  revolution.  Owing  to  the 
varying  lengths,  the  pipes  evenly  distribute  the  sand  in  the  vat. 
Water  and  slime,  floating  above  the  sand,  are  withdrawn  through 
four  vertical  gratings  set  equidistant  on  the  sides  of  the  tank.  One 
of  these  gratings  can  be  seen  at  the  edge  of  the  first  tank  in  Fig.  86. 
As  the  tank  fills  with  sand,  a  roller,  carrying  a  narrow  canvas 
curtain,  is  unrolled  to  the  level  of  the  deposited  material.  Another 
method  of  removing  slime  and  water  consists  in  entirely  filling  the 


182 


THE    METALLURGY 


tank  with  water,  and  letting  the  sand  drop  through  it,  the  surplus 
of  slime  and  water  overflowing  into  the  encircling  launder. 

The  overflow  from  d,  Fig.  85,  containing  $7.90  per  ton,  goes 
to  the  classifier  e,  which  removes  a  portion,  amounting  to  65%  of 
the  whole,  that  has  an  average  value  of  $9  per  ton.  This  clean  and 
finer  sand  is  distributed  in  the  tank  g' ',  while  the  overflow  from  the 
classifier  goes  by  the  pipe  A;  to  classifier  /.  The  overflow  from 
settling  tanks  g  and  -g'  also  flows  into  the  classifier  /.  The  object 
of  the  classifier  is  to  take  out  the  little  sand  that  escapes  settling 


Fig.    86.      TANKS  AND   BUTTERS  DISTRIBUTOR. 


previously.  The  amount  is  small  and  it  is  returned  by  the  pipe 
m  which  leads  back  to  the  receiving  tank  6,  and  caused  once  more 
to  go  through  the  system.  The  overflow  from  /,  consisting  only 
of  suspended  slime,  25%  the  weight  of  the  ore  crushed,  and  having 
a  value  of  $5  per  ton,  is  treated  with  additional  milk  of  lime  from 
an  automatic  feeder  n.  The  addition  of  the  lime  is  for  the  purpose 
of  hastening  the  settling  of  the  slime  that  flows  through  q  to 
settling  tanks,  while  the  clarified  supernatant  water  escapes  at  p, 
and  is  returned  to  the  battery. 

Assuming   that   the    tank   g}    which   holds   114   tons,    has   been 


OF    THE    COMMON    METALS.  183 

filled  with  the  mixture  of  coarse  sand  and  concentrate,  and  has 
been  drained,  it  receives  10  tons  of  weak  solution  (0.03%  KCN). 
After  this  the  ore  is  drained  and  transferred  to  the  lower  tank, 
this  part  of  the  treatment  occupying  five  days.  The  advantages 
of  making  the  transfer  are  two-fold:  first,  that  the  imperfectly 
packed  portion  becomes  mixed  with  the  other  sand,  and  second, 
that  the  material,  thus  moistened  with  cyanide  solution,  is  exposed 
to  the  oxygen  of  the  air  during  transfer.  In  the  lower  vat  the 
charge  receives  50  tons  of  strong  solution  (0.2%  KCN),  followed 
by  75  tons  of  medium,  and  65  tons  of  weak  solution.  It  is  then 
drained  and  discharged.  This  takes  twelve  days  more,  or  seventeen 
days  for  the  whole  treatment. 

The  fine  sand  in  /  requires  less  time  for  treatment  than  the 
coarse.  It  receives  10  tons  of  weak  solution  and  20  tons  of  medium 
solution.  It  is  then  treated  with  50  tons  of  strong  solution,  20  of 
medium,  and  20  of  weak.  The  time  is  three  days  for  the  upper, 
and  five  days  for  the  lower  vat. 

The  Blaisdell  excavator. — In  place  of  shoveling  the  contents 
of  the  upper  vat  into  the  lower  one,  an  operation  which  by  no  means 
breaks  up  the  lumpy  ore,  the  Blaisdell  system  for  transferring  and 
distributing  has  been  devised.  Fig.  87a  represents  the  excavator. 
It  consists  of  a  trussed  steel-bridge,  supporting  at  its  center  a 
vertical  steel  shaft  with  four  horizontal  arms  below,  called  excavator 
beams.  This  bridge  travels  on  rails  and  may  be  set  over  any 
vat.  The  central  vertical  portion  of  the  bridge  carries  guides  and 
screws  for  raising  or  lowering  the  shaft  and  arms.  The  arms 
carry  steel  harrow-disks. 

In  operation,  the  central  door  or  trap  at  the  bottom  of  the 
vat  is  opened  and  a  hole  is  quickly  dug  down  to  it  through  which 
the  sand  may  fall.  The  excavator  is  now  placed  in  position  and 
the  shaft  and  arms,  having  been  put  in  motion  by  an  independent 
motor,  the  steel  disks  pass  over  the  surface  of  the  sand,  cut  it, 
and  roll  it  toward  the  center  where  it  falls  down  through  the 
trap.  The  shaft  is  gradually  lowered  as  the  cutting  proceeds, 
much  as  an  auger  would  advance  in  boring  a  hole,  until  the  vat 
is  emptied.  The  arms  are  then  raised  and  the  excavator  moved  to 
another  vat. 

The  excavated  sand  from  the  settling  tank  falls  upon  a  belt- 
conveyor  which  carries  it  to  the  leaching  vat.  When  it  reaches 
the  vat  a  distributing  machine  or  distributor,  shown  in  Fig.  87b, 
that  consists  of  a  movable  steel  bridge  supporting  a  conveyor- 


Fig.    87a.      BLAISDELL   EXCAVATOR. 


Fig.    87b.      SAND   DISTRIBUTOR,   BLAISUELL,   SYSTEM. 


OF    THE    COMMON    METALS.  185 

belt,  takes  the  sand  from  a  tripper  and  drops  it  into  a  hopper  at 
the  center  of  the  bridge.  It  falls  from  the  hopper  upon  a  rapidly 
revolving  steel  plate  which  breaks  lumps  and  distributes  the  sand 
lightly  and  evenly  in  the  vat  in  an  ideal  condition  for  rapid  and 
uniform  leaching.  The  large  circular  band  or  ring,  suspended 
from  the  bridge,  is  to  prevent  throwing  the  sand  too  far. 

(a)  Decantation. — The  slime  coming  from  the  pipe  q,  Fig.  85 
is  collected  in  large  settling  tanks.     Each  tank  when  filled  is  cut 
out    from    the    flow,    and    allowed    to    settle.      After    settling,    the 
supernatant  water  is  drawn  off  by  the  decanting  device.  Fig.  76. 
This   removes   the   surface   water   continuously.     A  weak   solution 
(0.1%  KCN)   is  then  added  to  the  slime  in  the  settling  tank,  and 
agitation  effected  with  mechanical  stirrers,  or  by  transferring  the 
pulp  from  one  tank   to   another  with  a  centrifugal  pump.     After 
several  hours  the  solution  is  allowed  to  settle,  then  drawn  off  by 
decantation.    More  cyanide  is  then  added,  the  whole  stirred,  allowed 
to  settle,  and  half  the  solution  again  withdrawn.     The  operation  is 
repeated  several  times,  the  final  wash  being  of  water.    In  this  way 
75%   of  the  gold  in  the  slime  is  extracted. 

The  various  cyanide  solutions  mentioned  above  are  run  through 
the  zinc  precipitating  boxes,  the  zinc-shaving  having  first  been 
prepared  by  dipping  it  in  a  solution  of  lead  acetate.  The  zinc 
becomes  covered  with  a  film  of  lead  forming  a  zinc-lead  couple,  and 
is  more  active  in  depositing  gold  than  zinc  alone  would  be. 

The  objections  to  this  method  of  slime  treatment,  are :  the 
large  and  awkward  plant  required,  and  the  loss  due  to  the  impossi- 
bility of  saving  the  last  traces  of  dissolved  gold.  These  were  soon 
recognized  by  metallurgists  in  Western  Australia,  and  the  following 
method  was  devised  : 

(b)  The  filter-press  method  of  slime  treatment. — This  consists 
in  agitating  the  cyanide  containing  slime-pulp  (in  a  settling  tank) 
and   removing   this   solution   containing   the   gold,   by   forcing   the 
material  under  a  high  pressure  into   a  filter-press,   Fig.   78.     The 
solution  retained  in  the   slime-cake  is  displaced  by  forcing  dilute 
solution,    or   water,    through    the    press.      The    comparatively    dry 
compressed  cake  of  slime  containing  20  to  25%  water,  is  dropped 
into  a  car  beneath  and  trammed  to  the  dump.    The  filtered  solution 
goes  to  the  gold-solution  tank,  and  thence  through  the  zinc  boxes, 
to  the  proper  sump.     An  estimate  of  the  cost  of  this  kind  of  filter- 
pressing  is  38c.  per  ton. 

(c)  Suction  filtration. — With  the  exception  of  the  system  used 


186 


THE    METALLURGY 


at  the  Homestake  slime  plant,  the  comparatively  high  cost  of  filter- 
pressing  and  the  heavy  cost  of  installation  has  led  to  the  adoption, 
in  the  United  States,  .of  various  systems  of  vacuum  filter-pressing 
of  which  the  Butters  filter,  Fig.  88,  is  an  example. 

The  Butters  solution-filter. — The  white  rectangular  object,  at 
the  center,  in  the  background  of  the  illustration,  is  a  cluster  of 
four  filter-leaves  suspended  from  an  over-head  crawl.  These  have 
just  been  hoisted  out  of  the  tank  below.  A  single  leaf  is  10  ft. 
long  by  5  ft.  high.  The  upper  part  of  the  frame  of  the  leaf  is 


Fig.    88.      BUTTERS   FILTER  TANKS. 


a  wooden  bar,  sufficiently  long  to  span  the  tank  and  rest  upon  the 
sides.  The  remaining  three  edges  of  the  frame  are  of  iron  pipe. 
The  filter-leaf  consists  of  a  piece  of  cocoa-matting  cut  to  exactly 
fit  the  frame,  and  this  is  covered  on  each  side  by  a  sheet  of  canvas. 
These  three  sheets  of  material  are  sewed  together,  and  the  whole 
fitted  round  and  fastened  to  the  frame.  Five  vertical  strips  bolted 
on  the  outside  serve  to  stiffen  the  frame.  Fig.  88  gives  a  view  of 
two  60-frame  filter-tanks  which,  in  Fig.  89  and  90  are  marked  E,  E. 
Fig.  89  is  an  elevation  and  Fig.  90  a  plan  of  a  120-frame  Butters 
vacuum-filter  plant.  The  slime  is  in  cyanide  solution  several  hours, 
or  until  the  gold  is  dissolved.  For  convenience  the  pulp  is  collected 
in  a  tank  A,  called  the  pulp-storage  tank,  and  from  there  is  run, 
as  desired,  into  one  of  the  filter-tanks  E,  E,  filling  it  to  the  tops 
of  the  frames.  It  will  be  noticed  in  Fig.  88  that  the  pipes  of  the 


OF    THE    COMMON    METALS. 


187 


frames  are  connected  to  a  common  header.  This  header  leads  to 
a  vacuum  pump  N.  Upon  starting  the  pump,  the  solution  is  sucked 
through  the  filter-frames,  discharging  into  a  gold-solution  tank 
outside  the  building.  Pulp  is  run  into  the  box  at  intervals  to  keep 


Fig.  89.   GENERAL  ELEVATION  OF 
BUTTERS  FILTER  PLANT. 


Fig.  90.   GROUND  PLAN  OF  BUTTERS  FILTER  PLANT. 

the  filter-leaves  submerged.  The  slime  remains  as  a  cake  on  the 
outside  of  the  filter-leaves,  and  forms  a  deposit  sometimes  nearly 
an  inch  in  thickness. 

When  the  cake  is  formed  the  supply  of  pulp  is  shut  off,  and 
the  surplus  pulp  run  to  the  tank  C  and  pumped  back  to  A.    A  weak 


188  THE    METALLURGY 

cyanide  solution  from  tank  B  is  next  admitted  to  the  filter-tank, 
and  is  sucked  through  the  cake  to  displace  any  gold  solution 
remaining,  the  level  being  maintained  as  before.  The  surplus  wash- 
solution  is  pumped  back  to  B.  In  the  same  manner  the  cake  is 
washed  with  clear  water  to  displace  the  last  of  the  cyanide  solution. 
Both  washes  go  to  a  weak  gold-solution  tank  near  (}.  While  the 
tank  is  being  emptied  and  filled  and  the  cake  is  exposed  to  the 
air,  the  vacuum  must  be  reduced  to  5  in.  to  hold  the  cake  firmly  in 
place.  The  vacuum  pump  is  now  shut  off,  and  air  under  a  slight 
pressure  is  forced  into  the  leaves,  displacing  or  throwing  off  the 
cake,  which  drops  to  the  bottom  of  the  tank.  The  12-in.  quick- 
opening  discharge  gates,  6%  are  opened,  and  the  slime  is  washed  out 
into  a  waste-launder.  The  gates  are  again  closed,  and  the  filter 
is  ready  for  a  new  charge.  The  formation  of  the  slime-cake  adjusts 
itself  as  the  composition  of  the  material  varies  at  any  point,  the 
coarser  forming  a  thicker  layer  than  the  finer,  thus  maintaining 
the  even  permeability.  The  time  of  washing  is  15  minutes.  The 
operation  is  conducted  by  one  man.  A  precipitate  of  calcium 
carbonate  gradually  forms  upon  the  leaves,  decreasing  their 
permeability,  and  every  two  to  nine  months  this  is  dissolved  by 
dipping  the  leaves  in  tanks  J  which  contain  a  weak  solution  of 
hydrochloric  acid.  In  Fig.  88,  a  set  of  leaves  is  shown  in  the 
process  of  transferring  to  the  washing-tank.  The  operating  cost 
for  filtering,  including  labor,  power,  and  repairs,  in  a  large 
installation,  amounts  to  8c.  per  ton  of  slime  treated. 

43.     CYANIDATION    AT    THE    HOMESTAKE    MILL,     SOUTH 

DAKOTA. 

The  ore  is  a  garnetiferous  hornblende  schist  containing  7  to  8% 
pyrite  and  pyrrhotite.  It  is  crushed  with  eight  to  ten  times  its 
weight  of  water,  and  amalgamated,  using  inside  plates  for  the 
mortar,  as  well  as  outside  amalgamated  apron  plates. 

The  pulp  from  the  mill,  some  1300  tons  daily,  is  carried  by 
launder  to  a  cone-house,  Fig.  91,  to  a  set  of  twelve  settling  cones 
that  are  10  ft.  diam.  with  sides  having  an  angle  of  50  degrees. 
The  stream  is  distributed  evenly  to  all,  entering  each  at  the  center. 
Half  the  water  and  the  finest  slime  is  removed  by  the  overno\v. 
Some  fine  sand  is  carried  over  with  slime,  and  this  is  separated  by 
a  series  of  settling  tanks,  the  overflow  running  to  a  pond.  Here 
the  water  has  an  excellent  chance  to  settle,  and  when  nearly  clear, 
it  is  pumped  back  to  be  used  again  in  the  mill.  The  sediment 
accumulates  in  this  pond  and  finally  is  washed  out  with  a  hose  and 


OF    THE    COMMON    METALS.  189 

run  to  waste.  From  the  bottom  of  the  cones  is  withdrawn  a 
thickened  pulp  containing  the  sand  and  some  slime.  This  is  carried 
by  a  12-in.  pipe  on  a  2l/2%  grade  to  the  sand-plant,  Fig.  92  and  93. 

Classification  at  the  plant  is  effected  by  means  of  six  gravity 
settling  cones  6,  7  ft.  diam.  and  with  sides  of  a  50°  slope.  The 
underflow  from  each  of  these  goes  to  four  classifying  cones  c, 
24  in  all,  which  are  4  ft.  diam.,  and  with  steep  sides  of  70  degrees. 
These  are  provided  with  hydraulic  water,  as  in  the  lower  cone  of 
Fig.  99,  by  means  of  which  the  sand  settles  nearly  free  from  slime. 
The  slime  from  these  cones,  containing  90c.  to  $1  per  ton,  and 
amounting  to  40%  of  the  total  tailing,  was  formerly  wasted,  but 
more  recently  has  been  treated  by  filter-pressing,  at  the  slime- 
plant  to  be  described  later. 

The  sand-plant. — The  prepared  sand  contains  40.5%  coarse  par- 
ticles that  remain  on  a  100-mesh  screen;  30.8%  middles,  between 
100  and  200  mesh,  and  28.7%  fine  passing  200  mesh.  This  leaches 
at  the  rate  of  3  or  4  in.  per  hour.  Before  the  sand  enters  the 


Elevation  ofCo-neJfouse 


PJan  *f  Gynp-ffcvff 

Fig.    91.      CONE   CLASSIFIERS,   HOMESTAKE   MILL. 

leaching  vats  it  receives  a  stream  of  milk  of  lime  which  has  been 
prepared  by  being  stamped  in  a  five-stamp  battery  reserved  for 
the  purpose.  From  3  to  5  Ib.  of  the  lime  are  added  per  ton  of 
sand.  The  classified  pulp  and  lime  thus  mixed,  pass  to  a  Butters 
distributor,  see  Fig.  86,  which  can  be  transferred  from  one  vat  to 
another  by  an  overhead  trolley.  There  are  14  leaching-vats,  each 
44  ft.  diam.,  9  ft.  deep,  and  capable  of  holding  610  tons.  The  tank 
is  filled  with  wrater  and  the  sand  runs  in.  It  takes  11  hours  to 
charge  the  tank,  and  treatment  lasts  five  days.  When  the  tank 
is  filled  the  ore  is  drained  and  a  series  of  washes  of  the  stronger 
of  the  stock  solutions  (containing  0.14%  KCN)  is  run  in,  allowing 
each  wash  to  drain  below  the  top  of  the  ore  to  draw  in  air. 
The  effluent,  its  strength  reduced  to  0.10%,  is  run  to  the  two  weak- 


190 


THE    METALLURGY 


solution  precipitation  tanks  /  /,  Fig.  93,  each  26  ft.  diam.  by  19  ft. 
deep.  After  this,  the  weak  solution  is  brought  upon  the  charge  and 
retained  two  days  more.  The  solution  escaping  during  this  period 
is  run  to  the  two  strong-solution  collecting  tanks,  e  e.  .  This  is 
followed  by  a  water  wash  which  finally  reduces  the  unextracted 
gold  to  5  to  7c.  per  ton. 

The   charge   is  now   ready  for   sluicing   out.      This   is   done   by 
two  men  in  four  hours,  four  side  gates  and  one  bottom  gate  being 


OF  THE  COMMON  METALS.  191 

used  for  the  purpose.  The  8-oz.  duck  filter-cloth  underlaid  with 
another  of  cocoa  matting  is  washed  clean.  The  vat  is  then  filled 
with  water  and  is  ready  for  the  next  charging. 

Precipitation. — The  solution,  resulting  from  the  leaching  with 
strong  solution,  run  into  one  of  the  weak-solution  tanks  /  /, 
contains  $2  gold  per  ton.  When  the  tank  is  filled  the  stream  is 
turned  into  the  second  tank,  and  the  first  is  ready  for  precipitation. 
The  tank  holds  300  tons  of  solution  that  is  agitated  by  compressed 
air  admitted  from  a  pipe  across  the  bottom  of  the  tank  pierced  with 
numerous  small  holes  for  the  escape  of  the  air.  Sixty  pounds  of 
zinc  powder  in  the  form  of  an  emulsion  is  sprayed  in,  during  the 
agitation,  and  in  20  minutes  the  precipitation  of  the  gold  is  complete. 
A  duplex  pump  is  now  started  and  the  mixture  in  the  tank  is 
pumped  up  to  two  large  filter-presses,  Ti  and  i,  each  containing  24 
frames  36  in.  square.  The  filter-cloths  of  the  presses  soon  become 
coated  with  gold  precipitate  and  zinc-powder,  and  every  drop  of 
solution  passing  through,  comes  into  molecular  contact  with  the 
zinc  dust.  The  value  of  the  solution  is  reduced  from  $2  to  5  or 
lOc.  per  ton  (a  precipitation  of  97.5  to  95.0%)  and  is  then  termed 
barren  solution  and  passes  to  .the  weak-solution  storage  tanks  A*. 

The  weak-solution  wash  above  referred  to  fills  the  strong- 
solution  collecting-vats  e  e  and  is  strengthened  to  0.14%  KCN  and 
pumped  directly,  without  prior  precipitation,  to  the  strong- 
solution  storage  tank  /,  from  which  it  is  drawn  for  the  first  treat- 
ment of  the  sand. 

It  is  seen  that  the  strong  solution  of  one  day  becomes  the  weak 
solution  of  the  next,  and  that  all  the  gold  accumulates  in  the  weak- 
solution  precipitation  tanks.  The  strong  solution  has  an  approx- 
imately constant  value.  One-half  the  total  effluent  solution  is  pre- 
cipitated, the  other  half  having  a  nearly  constant  value  of  30  to 
50c.  per  ton. 

The  precipitate,  containing  gold  and  the  fine  zinc,  accumulates 
in  the  filter-presses,  that  are  run  a  month  without  opening.  In 
this  way  about  a  ton  of  precipitate,  worth  $50,000,  is  obtained. 
The  presses  are  opened,  the  precipitate  removed  and  sent  to  a 
lead-lined  mixing  tank  that  is  equipped  with  a  mechanical  agitator, 
a  hood,  and  an  exhaust  fan  for  removing  acid  fumes.  Here  it  is 
treated  with  dilute  hydrochloric  acid  to  dissolve  the  zinc,  agitated, 
and  settled.  The  supernatant  solution  is  then  drawn  off,  by  means 
a  montejus  or  pressure-tank  to  another  smaller  filter-press. 
Sulphuric  acid  is  next  added  in  the  mixing-tank  and  the  mixture 


192  THE    METALLURGY 

is  agitated  and  heated.  Upon  settling,  the  supernatant  acid  solution 
follows  the  first  one  through  the  press.  Wash-water  is  next  run 
into  the  tank  and  without  further  settling  is  run  to  the  press. 
Finally  the  precipitate  in  the  press  is  washed  with  clean  water. 
The  acid  solutions  and  the  wash-water  go  to  a  large  settling  tank 
that  acts  as  a  guard  to  recover  any  escaping  particles. 

The  acid-treated  precipitate  is  now  transferred  from  the  press 
to  a  large  steam-dryer,  where  a  part,  but  not  all,  of  the  moisture 
is  removed.  It  is  then  mixed  with  litharge,  borax,  silica,  and 
powdered  coke  and  sprinkled  with  lead  acetate  solution,  and 
briquetted  in  a  press  under  a  pressure  of  2  to  3  tons  per  square  inch. 
The  briquettes  are  dried,  charged  into  an  English  cupelling-furnace, 
see  Fig.  179,  and  quietly  fused,  suffering  no  loss.  The  lead,  as  in 
assaying,  absorbs  the  gold  and  sinks  to  the  bottom,  while  the 
slag  flows  away  as  it  forms.  The  'test'  gradually  fills  with  lead, 
which  is  oxidized  to  litharge  in  cupellation,  leaving  a  metal  975 
to  985  fine  in  gold  and  silver,  which  is  run  into  ingots.  The 
litharge  produced  is  reserved  to  be  added  to  the  precipitate  at 
the  next  clean-up. 

The  cupel-slag,  cupel-bottoms,  sweepings,  or  other  metal- 
bearing  cleanings  are  accumulated,  and  occasionally  run  through  a 
small  silver-lead  blast-furnace.  There  is  produced  a  slag  assaying 
less  than  $5  per  ton,  and  base-bullion  which  is  returned  to  the 
cupel-furnace  at  the  next  clean-up.  The  method  of  treating  the 
precipitate,  it  is  claimed,  is  practised  with  a  loss  of  only  0.1  per 
cent.  The  cost  of  treatment  per  ton  of  sand  is  26c.  This  has 
been  brought  into  comparison  with  South  African  costs  that,  before 
the  Boer  war,  varied  from  55  to  72c.  per  ton. 

The  slime-plant. — The  slime  pulp,  amounting  to  1600  tons  daily, 
has  an  average  value  of  91c.  per  ton.  It  contains  three  tons  of 
water  to  one  ton  of  solid,  and  is  carried  two  miles  by  a  12-in. 
pipe  at  a  grade  of  1.5%  to  the  slime-plant.  Here,  two  small  vats 
are  provided  for  slaking  lime.  The  content  is  drawn  to  a  screen- 
bottom  box  where  the  undissolved  lumps  separate.  The  box  over- 
flows into  an  agitator  from  which  the  milk  of  lime  continuously 
runs  into  the  main  slime-stream  at  the  rate  of  5  Ib.  of  lime  per 
ton  of  dry  slime.  Two  storage  vats,  26  ft.  diam.  and  24  ft.  deep, 
having  conical  bottoms  with  47°  sides,  receive  the  stream.  From 
the  bottom  of  these  storage  vats  the  slime-pulp  is  drawn 
continuously  through  a  10-in.  pipe  to  large  filter-presses  65  ft. 
below  to  obtain  a  pressure  of  30  Ib.  per  sq.  in.  The  10-in.  main 
extends  the  whole  length  of  the  press-building.  Between  each 


OF  THE  COMMON  METALS.  193 

pair  of  presses,  the  main  branches  into  10-in.  pipes,  which  in 
turn  send  two  4-in.  branches  to  each  press.  The  smaller  branches 
connect  to  a  4-in.  passage  or  channel  that  extends  along  the  center 
of  the  top  of  the  filter-press  frames.  From  the  channel  the  slime- 
pulp  flows  into  the  press.  There  are  92  frames  each  4  ft.  by  6  ft. 
and  4-in.  distance-frames  to  form  slime-cakes  4  in.  thick. 

The  press  having  been  filled,  a  0.1%  solution  of  KCN  is  run 
in,  entering  at  the  lower  corners  of  the  press  by  two  channels 
each  2.5.  in.  diam.  Compressed  air  is  admitted  at  the  upper  corners 
of  the  press  by  two  channels  each  2.5.  in.  diam.  This  is  followed 
by  a  0.04%  KCN  wash,  another  treatment  with  air,  and  again 
a  water-wash,  until  the  exhausted  slime  contains  no  more  than 
9  cents  gold  per  ton.  Along  the  center  of  the  bottom  of  the  frames 
is  a  continuous  6-in.  channel,  and  within  it  a  3-in.  pipe  extending 
the  length  of  the  press.  This  pipe  is  provided  with  92  nozzles 
1  in.  long  and  r>/32  in-  diam.  By  a  special  mechanism  a  revolving 
motion  through  200  degrees  is  given  to  the  pipe,  so  that  under  a 
water-pressure  of  50  Ib.  per  sq.  in.,  the  nozzles  play  against  the 
compact  slime-cakes,  removing  the  cake  completely  from  the 
compartments  in  45  minutes,  the  6-in.  channel  (generally  closed) 
being  then  opened  for  the  exit  of  the  slimes.  The  operation, 
exclusive  of  the  time  of  filling  and  emptying,  occupies  6  hours. 
It  will  be  noticed  that  the  solution  of  the  gold,  and  the  extraction 
of  the  slime,  is  done  entirely  at  one  operation  and  within  the 
press.  The  effluent  solutions  now  flow  to  the  strong  and  weak  gold 
precipitation-vats  where  the  gold  is  precipitated  by  means  of  zinc 
dust.  Zinc-dust  is  a  fine  powder  produced  in  retorting  zinc  ores 
and  contains  90%  Zn.  An  emulsion  of  this  zinc  in  water  is  added 
to  the  solution  in  a  conical-bottom  tank.  With  an  air-lift  pump 
a  thorough  mixture  of  the  zinc  with  the  solution  is  obtained. 

As  at  the  sand-vats,  the  solutions  are  pumped  to  the  storage 
solution-tanks  and  to  the  precipitate-presses,  the  treatment  at 
this  stage  being  the  same  as  that  already  described  as  taking  place 
at  the  sand-plant.  The  estimated  cost  of  this  plant  is  half  a  million 
dollars,  and  the  cost  of  treatment,  per  ton  of  slime,  is  25  cents 
including  all  items.  The  recovery  is  90  per  cent. 

44.     CYANIDATION  AT  EL  ORO,  MEXICO.  (THIRD  METHOD.) 

The  third  method  of  cyaniding  is  well  exemplified  in  the  practice 
at  El  Oro,  Mexico.  The  oxidized  ore,  a  hard  compact  quartz, 
contains  from  $6  to  $15  gold  and  3  to  5  oz.  silver  per  ton.  The 


194 


THE    METALLURGY 


gold  is  distributed  so  finely  through  the  quartz  that  seldom  can 
a  ' color'  or  visible  speck  be  discovered.  The  silver  is  partly  metallic, 
and  partly  in  the  form  of  sulphide,  arsenide,  and  antimonide.  Both 
metals  are  considered  to  be  deposited  between  the  faces  of  the 
elementary  quartz  crystals.  The  necessity  of  fine  grinding  to  unlock 
the  gold  and  silver  particles  is  well  illustrated  by  Fig  94. 

The  mine-ore  is  dumped  from  skips  at  the  shaft  into  ore-bins, 
coarsely  crushed,  and  conveyed  to  the  mill  by  cars.  It  is  stamped 
through  a  25-mesh  screen,  then  goes  over  amalgamated  plates  which 
catch  from  13  to  18%  of  the  gold  and  1  to  5%  of  the  silver.  The 


100% 


Fig.  94.  RELATION  BETWEEN  MESH  AND  EXTRACTION. 

pulp  from  the  plates  is  re-ground  through  tube-mills,  then  flows 
to  the  cyanide  plant  where  it  is   separated  into  sand   and  slime, 
these  being  treated  separately.     The  daily  duty  per  stamp   is  4.7 
tons,  and  the  ratio  of  water  to  ore  is  as  nine  or  ten  to  one. 
.    Fig.  95  is  a  flow-sheet  showing  the  method  of  treatment. 

The  ore  pulp  from  the  plates  passes  to  two  cones  4.5  ft.  diam. 
beneath  which,  in  series,  are  two  2-ft.  cones  called  'pulp  thickeners.' 
The  overflow  from  all  the  cones  gives  a  product  containing  81% 
slime,  while  the  underflow  or  spigot  discharge,  a  sand  still  retaining 
9%  slime,  passes  to  the  tube-mill  No.  5,  where  half  of  it  is  ground 
to  a  slime. 


OF    THE    COMMON    METALS. 


195 


Fig.  96  represents  a  view  and  Fig.  97  a  longitudinal  section 
of  a  tube-mill.  It  consists  of  a  steel  shell,  like  a  boiler-shell,  5  ft. 
diam.  by  23  ft.  long,  making  27  revolutions  per  minute,  and  having 
a  daily  capacity  of  130  to  150  tons.  It  is  lined  with  silex  blocks 


OF    THE    COMMON    METALS. 


197 


(a  compact  quartz)  about  the  size  of  ordinary  brick,  and  in 
operating  is  filled  with  hard  quartz  pebbles  the  size  of  the  fist. 
As  the  barrel  revolves,  these  pebbles,  dropping  back  upon  one 
another,  finely  grind  any  material  fed  to  the  mill.  The  ore  is 
fed  in  from  a  hopper  placed  at  one  end.  The  ground  pulp  escapes 
through  the  perforated  false  head  at  the  other  end,  and  overflows 
through  the  other  trunnion. 


198  THE   METALLURGY 

Referring  to  Fig.  95,  the  discharge  joins  the  overflow  of  the 
cones  and  goes  to  spitzkasten  No.  1,  which  takes  out  the  insufficiently 
ground  sand.  This  is  raised  by  the  elevator  wheel  and  sent  to 
return  cone  No.  5  (the  purpose  of  which  is  to  remove  the  unground 
sand),  then  goes  to  tube-mills  No.  3  and  4  for  re-grinding  and 
finally  back  to  spitzkasten  No.  1,  while  the  fine-sand  overflow  of 
return  cone  No.  5  passes  to  one  of  the  nine  'sand-receivers'  over  a 
Butters  distributor.  The  overflow  of  these  sand-receivers  goes  on 
to  spitzkasten  No.  2,  which  takes  out  a  little  fine  sand  which 
returns  to  the  elevator  wheel,  and  an  overflow  that  joins  the  over- 
flow from  spitzkasten  No.  1  and  passes  on  to  the  'slime  tanks.' 

Each  sand  receiver,  when  filled,  is  drained  24  hours,  and  with- 
out further  preliminary  treatment,  is  discharged  by  a  Blaisdell 
excavator,  see  Fig.  8jTa,  upon  a  belt-conveyor  that  takes  it  to  one 
of  the  12  treatment  tanks  to  be  charged  by  a  Blaisdell  distributor, 
see  Fig.  87.  In  these  tanks  the  sand  is  regularly  leached. 

The  treatment  consists  in  leaching  the  sand  with  a  certain 
number  of  strong  and  medium  solutions  or  washes,  given  alternately, 
and  followed  by  a  number  of  weak-solution  washes.  By  these, 
the  gold  in  the  tailing  is  reduced  to  20c.  per  ton,  and  the  total 
time  of  leaching  is  90  hours.  The  use  of  the  sand-receivers  to  do 
the  first  part  of  the  leaching  has  been  discontinued  because  of  the 
loss  of  time  and  cyanide  solution,  and  because  the  tank  becomes 
tightly  packed  and  uneven.  The  aeration  effected  by  excavating 
and  redistributing  the  sand,  has  a  great  influence  on  the  rate  of 
extraction.  This  is  helped  by  the  speed  at  which  the  washes  follow 
one  another.  The  gold  is  more  speedily  extracted  than  the  silver. 
Cyanide  consumption  is  also  rapid  at  first.  Upon  the  completion 
of  the  treatment  the  tailing  is  removed  by  another  Blaisdell 
excavator  and  by  a  troughed  conveying-belt  that  takes  it  to  the 
dump. 

The  slime-pulp,  overflowing  from  spitzkasten  No.  2  is  composed 
of  1  part  of  slime  to  12  of  water.  To  cause  the  slime  to  settle 
readily,  caustic  lime  is  added  to  the  mill-pulp  (just  before  it 
enters  the  tube-mills)  at  the  rate  of  12  Ib.  lime  per  ton  of  ore.  The 
slime-pulp  flows  continuously  into  one  of  the  slime-tanks,  75  tons 
of  slime  settling  in  the  bottom  of  the  tank,  while  the  clear 
supernatant  portion  overflows  to  the  'mill-water  sumps,'  whence  it 
is  pumped  back  for  use  in  the  mill.  When  a  charge  has  been 
completed  the  slime  is  allowed  to  settle  8  hours,  and  the  clear 
water  decanted.  The  slime  is  now  stirred  by  means  of  the 
mechanical  agitator  of  the  slime-tank.  In  ten  minutes  thereafter  a 


OF    THE    COMMON    METALS.  199 

centrifugal  pump  is  started  by  which  the  pulp  is  elevated  and 
thrown  back  into  the  tank.  The  pulp  having  been  thoroughly  mixed, 
a  specified  amount  of  0.05%  cyanide  solution  as  well  as  lead  acetate 
(0.4  Ib.  per  ton  of  slime)  is  added,  and  the  whole  agitated  7 
hours.  The  tank  is  filled  with  fresh  cyanide  solution,  the  stirring 
is  stopped,  the  content  allowed  to  settle,  and  the  solution  decanted 
as  closely  as  possible.  Four  additional  washes  are  given  as  rapidly 
in  succession  as  the  settling  of  the  slime  will  allow.  The  residual 
slime  is  then  pumped  into  a  slime-settler,  is  gradually  settled,  then 
decanted,  and  the  discharge  slime  run  to  waste. 

It  will  be  seen  that  there  are  two  of  the  gold-solution  tanks, 
the  'strong'  and  the  'medium,'  which  take  their  supply  from  the 
sand-tanks,  while  a  group  of  three  (the  weak-solution  tanks)  take 
their  supply,  not  only  from  these,  but  from  the  slime-tanks  as  well. 
The  final  wash,  which  is  decanted  from  the  slime-settlers,  contains 
but  little  gold  and  need  not  go  through  the  zinc-boxes  but  instead, 
to  the  weak-solution  sump-tanks.  To  aid  precipitation,  at  the 
strong  and  medium  zinc-boxes,  a  regulated  flow  of  2.5%  KCN  is 
added  to  the  inflowing  solution  from  the  sand-plant.  After  this 
has  passed  the  boxes,  it  is  raised  to  its  full  strength  by  the  addition 
of  a  concentrated  solution  of  KCN  contained  in  a  'dissolution  tank' 
near  the  sumps.  When  pumped  to  the  slime-tanks  it  is  there 
strengthened  by  letting  it  run  over  a  sack  containing  solid  KCN 
salt.  At  the  sides  of  both  the  slime  tanks  and  the  slime-settlers 
are  swinging  decanting  pipes,  which  withdraw  the  upper  portion 
of  the  clear  supernatant  solution. 

45.     CYANIDATION  IN  SOUTH  DAKOTA  (FOURTH  METHOD). 

An  example  of  this  method  is  found  at  the  Maitland  mill, 
Maitland,  South  Dakota.  The  ore  is  close-grained,  hard,  and  in 
the  main  oxidized,  through  carrying  a  little  pyrrhotite  and  pyrite. 
It  averages  75%  SiO2,  11%  Fe,  1.2%  S,  and  contains  $9  gold  and 
0.5  oz.  silver  per  ton.  The  gold  occurs  evenly  distributed  and 
finely  divided,  so  that  cyanidation  is  necessary  to  recovery. 

Fig.  98  is  a  plan  of  the  mill  and  indicates  the  operation.  The 
ore  from  the  mine  is  dumped  into  a  flat-bottom  bin,  the  idea  being 
to  permit  the  ore  to  form  its  own  slope,  thus  avoiding  wear  upon 
the  bottom  of  the  bin.  It  is  coarsely  crushed  by  a  large  24  by  13 
in.  Blake  crusher  to  1.5-in.  size,  and  is  elevated  by  a  belt  elevator 
to  the  battery  bin,  a  sample  amounting  to  2%  of  the  ore  being 
taken  at  the  elevator  head.  The  ore  from  the  battery  storage-bins 
is  fed  by  automatic  feeders  as  shown  in  Fig.  45,  to  the  40-stamp  mill 


200 


THE    METALLURGY 


of  910-lb.  stamps,  where  it  is  wet-crushed  in  cyanide  solution 
through  a  26-mesh  screen  at  the  rate  of  3  tons  daily  per  head  or  120 
tons  in  all.  The  battery  solution  is  kept  at  a  strength  of  0.06% 


SAMPLE  ROOM 


r 

Sliiue      Vyfsiime  ' 

to  Jistril.utor  over  vata 1_/V_L  Cle.in  sand  to  distributor  over  VMS  i 


1 ._  .      J_qbld_  '^solution  _C°L.LE,CJIN 


\ 

\ 


/BATTERY)    "I           (  GOLD    V  Cyanide  tot 

V         fiOL          /  \         TA     \t        I  

VACUUM!/    .wash  water 

net  to    FII>L^I      J  Gold  sol. 


Fig,    98.      PLAN  OF  OPERATION  AT   THE   MAITLAND  MILL 


OF    THE    COMMON    METALS. 


201 


cyanide,  and  has  0.05%  protective  alkalinity.  Four  to  five  tons 
of  solution  are  used  per  ton  of  ore.  In  crushing,  4  to  6  Ib.  quick- 
lime is  added  in  the  battery  to  assist  later  in  the  precipitation  of 
the  slime. 

The  proper  separation  of  sand  from  slime  is  important.  It  is 
performed  in  the  following  way :  The  battery-discharge  is  elevated 
by  Frenier  sand-pumps  to  the  distributor-box  of  the  cone  system 
as  shown  in  Fig.  99.  The  feed  in  this  box  is  divided  evenly  between 


discharge 
Distributing  box  fo  sand  tanks 

Fig.    99.      SINGLE-CONE   CLASSIFIERS. 

two  simple  settling  cones  50  in.  diam.  using  no  hydraulic  water, 
so  that  as  the  sand  settles  and  discharges  at  the  bottom  of  the 
cones,  it  still  carries  slime.  This  spigot  discharge  is  received  in 
a  launder  which  delivers  it  to  a  50-in.  hydraulic  cone-classifier 
having  a  hydraulic  supply,  not  of  water  but  of  battery  solution. 
Thus  a  clean  sand  is  separated  at  the  spigot  discharge  that  contains 
no  more  than  1  to  5%  slime,  while  the  slime  with  much  of  the 
solution  joins  the  overflow  from  the  first  cones. 

Treatment    of    the    sand, — The    clean    sand    from    the    spigot 
discharge  of  the  lower  cone,  amounting  to  half  the  total  weight  of 


202  THE    METALLURGY 

ore,  and  containing  2x/2  to  3  parts  solution  to  one  of  sand,  Hows 
by  launder  to  the  Butters  distributor  and  fills  the  tanks  evenly. 

The  six  '  sand- vats, '  holding  140  tons  each  are  30  ft.  diam.  by 
6  ft.  deep,  with  lattice  bottoms  as  shown  in  Fig.  68.  The  lattice 
is  covered  with  cocoa-matting  and  the  matting  with  8-oz.  duck. 
The  filter  lasts  10  months.  A  tank  is  filled  in  60  hours.  A  system 
of  dry  filling  is  used,  the  vat  not  being  filled  first  with  water.  All 
solution  coming  in  with  the  sand  is  allowed  to  drain  through  the 
filter-bottom.  This  gives  a  more  porous  and  easily  leached 
product,  since  the  slime  is  evenly  distributed  through  the  sand 
as  it  could  not  be  in  a  vat  first  filled  with  solution.  The  average 
weight  of  a  cubic  foot  of  the  sand  is  93  pounds. 

The  vat,  having  been  filled,  is  leveled  with  a  stream  of  solution 
from  a  hose  under  low  head.  Battery  solution  is  then  run  on  for 
a  period  of  10  days.  A  small  amount  of  slime  in  the  solution  forms 
a  slight  coating  on  the  sand,  and  to  insure  satisfactory  leaching, 
this  occasionally  is  lightly  raked  to  keep  it  pulverous.  The  battery- 
solution  is  followed  by  barren  solution  for  six  days  more.  This 
is  allowed  to  drain  and  is  followed  by  a  water-wash,  amounting 
to  10%  by  weight  of  the  sand.  To  treat  a  charge  of  140  tons 
of  sand  requires  900  tons  of  battery  solution  (exclusive  of  that 
which  has  passed  through  the  filter  filling)  and  450  tons  of  barren 
solution.  This  large  quantity  of  solution  (ten  parts  of  water  to 
one  of  sand)  and  prolonged  treatment  (16  days)  is  by  no  means 
excessive,  since  experience  shows  that  much  weak  solution  must 
be  constantly  percolating  through  the  charge. 

The  exhausted  sand  is  sluiced  through  a  launder  of  8%  grade 
with  100  to  150  tons  of  water.  The  tank  is  then  ready  for  another 
charge. 

Treatment  of  the  slime. — Referring  to  Fig.  98,  we  note  that 
the  overflow  from  the  upper  and  lower  cones  is  united  into  a  single 
flow,  which  goes  by  launder  to  one  of  the  slime  vats,  No.  4  or  5, 
called  'loading-vats.'  Each  loading-vat  has  a  central  partition 
extending  to  30  in.  from  the  bottom.  When  the  vat  is  filled  the 
stream  of  slime-pulp  continues  to  enter  the  vat  quietly  on  one 
side,  the  solid  portion  settling  out,  while  the  nearly  clear  solution 
escapes  over  the  rim  of  the  tank  at  the  opposite  side.  The  vats 
are  24  ft.  diam.  by  12  ft.  deep  and  hold  150  tons,  and  receive 
during  12  hours  300  tons  of  slime-pulp  containing  12  tons  of  solution 
to  one  of  slime.  Thus  150  tons  of  slime  is  decanted  and  30  tons 
is  retained  in  the  vat.  It  is  allowed  to  settle  and  is  decanted  as 
closely  as  possible,  then  pumped  by  a  centrifugal  pump  to  slime 


OF   THE    COMMON   METALS.  203 

vat  No.  1,  barren  solution  being  also  added.  Meanwhile  the  other 
loading-vat  is  filled,  settled,  and  transferred  to  slime  vat  No.  2  with 
barren  solution.  Both  vats,  No.  1  and  2,  are  now  decanted  and  the 
contents  combined  and  transferred  to  vat  No.  3,  making  a  fuil 
charge  of  60  tons  of  slime.  Two  more  transfers,  with  barren 
solution  added  each  time,  are  made  to  slime-vats  No.  6  and  Y,  and 
finally  "a  transfer  to  slime-vat  No.  8  where  the  charge  receives  a 
final  water-wash.  After  each  transfer  and  dilution,  several  hours 
of  agitation  are  given  by  pumping  from  the  bottom  of  the  vat  and 
discharging  into  the  top.  After  decanting  the  wash-water,  the 
slime  contains  55  to  60%  moisture.  The  layer  of  thin  slime  is 
drawn  back  on  the  next  charge,  leaving  a  dry  slime  with  but 
47%  moisture.  This  is  sluiced  out  and  run  to  waste. 

Course  of  the  solutions. — The  rich  gold-bearing  solutions  from 
the  sand-vats,  the  only  ones  that  run  through  zinc-boxes,  are 
received  in  a  collecting-box  and  pass  to  the  gold  tank  where  they 
are  united  and  brought  to  standard  strength  by  the  addition  of 
fresh  KCN.  This  promotes  precipitation  at  the  zinc-boxes.  The 
poor  solutions  from  the  sand-vats,  and  the  decanted  and  overflow 
solutions  from  the  slime,  still  called  battery  solutions,  flow  to 
another  collecting  box  and  thence  to  the  battery  sump.  The  solution 
in  the  sump  is  pumped  back  to  the  two  battery-solution  stock-tanks 
for  use  again.  It  is  not  yet  sufficiently  rich  in  gold  to  justify 
passing  it  through  the  zinc-boxes. 

It  is  thus  seen  that  the  cyanide  solution  makes  a  closed  circuit, 
the  only  part  escaping  being  that  in  the  exhausted  sand  and  slime. 
The  consumption  of  potassium  cyanide  is  but  0.84  Ib.  per  ton  of  ore 
treated. 

Precipitation. — From  the  gold-tank  the  solution  passes  in  a 
regulated  flowr  to  the  five  zinc-boxes  each  of  224  cu.  ft.  capacity, 
having  eight  compartments.  Here  the  gold  is  precipitated.  The 
barren  solution  passing  on,  enters  the  barren-solution  sump  whence 
it  is  pumped  back  to  the  barren-solution  stock-tank  at  the  highest 
part  of  the  mill.  Of  this  solution,  amounting  to  some  400  tons 
daily,  25%  is  used  for  the  treatment  of  sand  and  the  remainder 
for  the  slime.  The  consumption  of  zinc  averages  1.33  Ib.  per  ton 
of  ore  treated. 

Cleaning  up. — A  clean-up,  varying  somewhat  in  detail  from  that 
described  under  the  *  first  method,'  is  made  twice  monthly.  To  clean 
up  a  box,  the  flow  from  the  gold-solution  tank  is  shut  off  and  water 
run  in  15  minutes  to  displace  the  solution  in  the  shaving.  The 
zinc  in  the  first  compartment  is  shaken  up  and  down  to  remove 


204  THE    METALLURGY 

adhering  precipitate  and  set  aside.  The  water  is  bailed  close  to 
the  precipitate  in  the  compartment  into  the  next  compartment.  The 
plug  at  the  bottom  of  the  compartment  is  then  withdrawn  and  the 
remaining  water,  with  some  precipitate,  Hows  through  a  launder  to 
the  acid-tank.  The  screen-bottom  is  taken  out,  and  the  remaining 
precipitate  is  removed,  placed  in  a  tub,  and  carried  to  the  acid- 
tank.  The  compartment  is  washed  with  a  hose,  the  water  going  to 
the  acid-tank.  The  plug  and  screen  is  replaced,  and  the  zinc  shaving 
put  back,  with  shaving  from  the  second  compartment,  to  completely 
fill  the  first.  The  compartments  are  cleaned  in  succession,  the  zinc- 
shaving  being  moved  toward  the  head  until  all  compartments  HIV 
filled.  In  filling  the  first  compartment  a  slight  flow  of  solution  is 
started  to  avoid  undue  exposure  to  the  air,  and  the  water  from  the 
last  compartment  is  poured  into  an  adjoining  zinc-box.  The  acid- 
tank  thus  receives  all  the  precipitate  and  some  of  the  wash-water. 
One  man  cleans  up  the  five  boxes  in  12  hours. 

The  precipitate  is  allowed  to  settle  in  the  acid-tank,  then  the 
clear  supernatant  solution  is  closely  syphoned  into  a  waste  tank 
that  holds  25  tons  of  water,  and  shown  in  Fig.  98.  Concentrated 
sulphuric  acid  is  next  added  to  the  acid  tank  and  the  mixture  is 
stirred  in  by  hand  to  avoid  boiling  over.  The  treatment  lasts  an 
hour.  To  avoid  fume  a  well  ventilated  portion  of  the  precipitating 
room  is  chosen  for  the  acid-tank.  Acid  treatment  being  now 
complete,  the  tank  is  partly  filled  with  water,  the  vacuum  filter 
started,  and  the  entire  content  of  the  acid-tank  filtered  without 
further  washing.  The  filtrate  goes  to  the  waste  tank. 

The  precipitate  is  blown  from  the  filter-frame,  collected  in  the 
hopper-shaped  bottom  of  the  filter  vat,  withdrawn  in  pans,  and 
taken  to  the  melting-room  and  dried  in  a  muffle  furnace.  To  10 
parts  of  the  dry  precipitate  are  added  and  mixed  4  parts  soda, 
1  part  borax,  1.5  parts  sand  and  0.2  parts  fluorspar.  The  mixture 
is  melted  quickly  in  a  No.  200  graphite  crucible,  in  a  wind-furnace 
having  a  forced  or  under-grate  blast.  The  content  of  the  crucible 
with  the  slag  is  poured  into  molds  and  the  slag  removed  upon 
cooling. 

This  slag  contains  matte,  carrying  as  much  as  $10  gold  per 
pound.  It  is  accordingly  remelted  with  sand,  flux,  and  10%  by 
weight  of  scrap-iron.  The  fusion  gives  a  coppery  bullion  700  fine 
in  silver  and  80  fine  in  gold,  and  a  slag  that  is  shipped  to  a  smelting 
works.  The  ingot,  weighing  200  ounces,  is  treated  with  hot  concen- 
trated nitric  acid  in  a  porcelain-lined  kettle  and  gives  a  residue 
containing  50%  Au  and  25%  Ag.  This  residue  is  added  to  the 
zinc-box  precipitate  of  the  next  clean-up,  while  the  acid  solution, 


OF    THE    COMMON    METALS.  205 

containing  most  of  the  silver,  is  treated  with  soda  which  precipitates 
the  silver  carbonate.  This  is  roasted  at  a  low  red  heat,  and  yields 
silver  which  is  easily  fluxed  and  melted  in  a  crucible  and  cast  in 
the  form  of  a  bar. 

The  united  solutions  from  the  acid  tank  and  the  suction  filter 
are  retained  in  the  waste  sump.  They  contain  $2  gold  per  ton. 
The  solution  is  treated  writh  fine  zinc  to  precipitate  the  gold; 
sulphuric  acid  is  added  and  the  mixture  well  stirred  to  remove  the 
excess  of  zinc.  Some  90%  of  the  gold  is  thus  precipitated.  The 
sweepings  around  the  zinc-boxes  are  thrown  into  the  sump  and 
their  content  recovered  once  in  six  months. 

The  extraction  at  the  Maitland  mill  in  1901  amounted  to  75.5% 
of  the  gold  and  44.3%  of  the  silver.  The  total  cost,  including 
general  expense,  but  not  depreciation  of  the  plant  was  $1.61  per 
ton  of  ore  treated. 

46.     CYANIDATION     OF     CRIPPLE     GREEK     ORE      (FIFTH 

METHOD). 

The  ores  of  the  Cripple  Creek  district,  Colorado,  are  phonolites 
and  are  of  two  kinds:  first,  the  weathered  surface  ore,  having 
little  or  no  tellurium  and  containing  free,  finely  disseminated  gold; 
and  second,  the  deeper-lying  ores  containing  gold  combined  with 
tellurium  as  syivanite  and  calaverite,  with  some  nearly  barren  pyrite. 
The  surface  ores  are  treated  raw.  The  telluride  ores  receive  an 
oxidizing  roast  before  cyanide  treatment.  The  treatment  of  the 
telluride  ores  we  are  now  to  consider. 

Metallic  Extraction  Co.,  Cyanide,  Colorado. — The  ore,  coarsely 
crushed  to  %-in.  size,  is  stored  in  bins,  then  supplied  to  a  revolving 
dryer,  and  after  drying  crushed  to  30-mesh  size,  as  described  in 
connection  with  Fig.  25.  The  product  is  elevated  to  storage  hoppers 
from  which  a  regulated  feed  is  delivered  to  the  roasting-furnaces. 
The  roasting  not  only  breaks  the  combination  of  gold  and  tellurium, 
but  makes  the  ore  porous,  and  accessible  to  the  solution,  and  causes 
it  to  be  more  readily  leached.  The  roasting  is  done  in  one  of  the 
mechanical  roasters  of  the  Argall  or  the  Ropp  straight-line  type. 
The  cooling  apparatus,  which  receives  the  hot  ore  from  the  roasting 
furnace,  is  a  flat-bottomed  trough,  water-cooled  by  pipes  beneath. 
The  ore  (first  mixed  with  10  Ib.  quicklime  per  ton)  is  moved  along 
by  scrapers  which  deliver  it  to  a  storage  pit  at  the  end  of  the 
trough.  At  the  discharge  end,  the  ore  is  sprinkled,  before  leaving 
the  trough,  with  the  cyanide  solution.  This  cools  it,  prevents 
dusting,  and  starts  dissolution. 


206  THE    METALLURGY 

The  drying  and  roasting-furnaces,  as  well  as  the  housing  of  the 
dry  crushing  machines,  are  connected  by  flues  to  two  fan  blowers, 
each  capable  of  delivering  10,000  cu.  ft.  air  per  minute.  These 
fans  deliver  the  dust  to  a  bag-house,  described  under  the  'Bartlett 
Process.'  Dust,  thus  recovered,  is  about  twice  the  value  of  the  ore 
from  which  it  arises.  It  is  briquetted  in  a  White  briquetting-press, 
Fig.  164,  and  smelted. 

The  leaching  tanks  are  50  ft.  diam.  by  6  ft.  deep,  constructed  as 
shown  in  Fig.  70  and  71,  each  holding  400  tons  of  ore.  In  operation, 
the  tank  is  filled  with  ore  from  two-wheeled  buggies,  and  0.7%  KCN 
solution  is  admitted  below  under  pressure,  so  that  the  ore  and 
solution  rising  at  the  same  rate  fill  the  tank  in  50  hours.  The 
solution  stands  on  the  ore  24  hours,  then  more  is  added  and  drawn 
off.  (This  practice  is  inferior  to  that  of  actually  percolating  the 
ore.)  The  operation,  repeated  several  times,  reduces  the  gold 
content  to  one  third  of  the  original  in  five  days. 

A  weak  solution  of  0.25%  KCN  is  used  in  the  same  way  for 
three  days.  The  charge  is  then  given  several  water  washes,  drained, 
and  finally  washed  out  with  water  under  high  pressure  from  a  hose 
into  a  launder,  the  washing  and  discharging  taking  two  days  more. 
At  Cyanide,  where  the  plant  is  situated  on  a  level  site,  the  tailing 
is  elevated  by  a  centrifugal  pump  to  a  launder  on  a  high  trestle, 
discharged  to  a  dump,  and  impounded  behind  a  low  dam.  In 
the  launder  is  a  series  of  transverse  riffles  behind  which  any  coarse 
particles  of  gold,  produced  by  roasting,  are  caught.  These 
particles  are  likely  to  be  found  in  roasted  telluride  ore,  in 
consequence  of  the  condition  of  the  gold  when  first  released  from 
combination  with  tellurium  that  causes  it  to  form  drops,  shot,  or 
other  coarse  particles.  As  an  alternative  to  washing  out  the  tailing, 
it  may  be  shoveled  through  side-discharge  doors,  Fig.  72,  and 
trammed  away,  a  more  expensive  method. 

The  cycle  of  treatment  as  above  detailed  was  12%  days. 

The  value  of  the  original  ore,  charged  to  the  vat,  may  be  assumed 
to  be  0.90  oz.  Au  and  0.5  oz.  Ag  per  ton,  while  the  tailing  retains 
0.06  oz.  Au  and  0.2  oz.  Ag  per  ton.  The  extraction  accordingly 
is  93.33%  for  the  gold  and  60%  for  the  silver  while  the  loss  in 
cyanide  is  1.75  Ib.  per  ton  of  ore  treated.  The  assay  of  the  tailing 
shows  as  good  extraction  from  the  grains  of  40-mesh  size  as  from 
those  passing  a  200-mesh  screen,  and  indicates  how  effective  roasting 
is  in  preparation  for  subsequent  leaching. 

Each  kind  of  solution  passes  to  a  separate  gold-solution  tank 
and  is  run  in  regulated  flow  through  separate  zinc  boxes  to  its  own 


OF    THE    COMMON    METALS. 


207 


sump  to  be  strengthened  for  re-use.  The  zinc  consumed  amounts  to 
0.9  Ib.  per  ton  of  ore  treated.  The  precipitation  and  the  details 
of  subsequent  operation  are  described  under  the  *  first  method.'  The 
cost  of  treatment  of  Cripple  Creek  ore  is  $2  to  $2.50  per  ton 
including  roasting,  when  performed  on  the  large  scale  indicated. 
The  Golden  Cycle  Mill,  Colorado  City,  Colorado.— The  ore  from 
Cripple  Creek,  treated  at  the  mill,  is  coarsely  crushed  to  2-in.  size 


Fig.   100.     MONADNOCK    (CHILEAN)    MILL. 

in  rock-breakers  and  passed  through  coarse-crushing  rolls  which 
reduce  it  to  one  inch  after  which  a  sample  is  removed  automatically. 
The  ore  is  taken  by  a  belt-conveyor  to  storage  bins  holding  8000 
tons.  This  collects  the  ore  into  large  lots  and  insures  greater 
uniformity  of  treatment.  The  ore  is  removed  from  the  storage  bins 
on  a  troughed  belt-conveyor  and  is  delivered  to  intermediate  rolls 
able  to  crush  without  preliminiary  drying*  to  ^-in.  size.  This 
rather  coarse  ore  is  delivered  to  Edwards  roasters,  see  Fig.  38  and 
39.  It  has  been  found  that  roasting  can  be  thoroughly  done  with 


208  THE    METALLURGY 

ore  no  finer  than  this.  The  telluride  of  gold,  where  segregated 
in  pieces,  forms  particles  of  gold  in  roasting  that  are  too  large 
to  be  dissolved  in  the  potassium  cyanide.  The  roasted  ore,  after 
cooling,  is  ground  to  60  mesh  in  cyanide  solution  in  Chilean  mills, 
see  Fig.  100.  The  bed  of  the  mill  carries  a  shallow  pool  of  mercury, 
and  large  particles  of  gold  are  amalgamated  and  retained  here. 
It  has  been  found,  that  the  extraction  is  quite  as  good  when  the 
ore  is  ground  to  this  mesh  as  when  ground  to  200  mesh.  The  mill 
in  the  illustration  is  shown  with  the  surrounding  screens  removed 
to  expose  the  rollers.  The  rollers  travel  upon  a  ring-die,  the 
crushing  being  done  between  the  die  and  the  roller.  The  ore  is 
fed  into  the  hopper  near  the  top  and  enters  through  pipes  in  front 
of  the  rollers.  A  scraper  in  front  of  each  roller  deflects  the  pulp 
in  front,  but  is  so  set  as  to  escape  the  mercury  pool  inside  the  ring- 
die.  The  ore,  when  ground  sufficiently  fine,  splashes  out  through 
the  screens,  and  is  conducted  by  the  circular  cast-iron  launder  that 
surrounds  the  mill  to  the  point  of  discharge. 

The  resulting  pulp  is  elevated  to  classifiers,  and  separated 
as  usual  into  sand  and  slime,  the  sand  being  conducted  directly 
to  the  leaching  tank,  while  the  separated  slime,  after  four  hours 
agitation  in  a  tall  conical-bottom  tank,  is  drawn  to  Argall  vacuum 
filter-presses  of  the  Moore  type. 

47.     ACTION  OF  COPPER  IN  CYANIDING. 

Copper  in  ore  is  generally  understood  to  be  a  serious  detriment 
to  successful  cyanidation,  since  it  passes  into  solution  and  consumes 
cyanide.  The  difficulty  can  be  overcome,  \vhere  little  copper  is 
present,  by  running  strong  cyanide  solution  into  the  stream  of  gold- 
solution  as  it  enters  the  zinc-box.  This  strengthens  the  entering 
solution  and  prevents  the  precipitation  of  copper.  The  copper  that 
enters  the  solution  consumes  potassium  cyanide. 

The  Hunt  ammonium  cyanide  process. — This  is  a  simple  process 
dependent  on  the  fact  that  ammonium  chloride  added  to  potassium 
cyanide  solution  has  a  protective  influence  upon  the  cyanide  and 
a  solvent  action  upon  copper. 

At  Dale,  San  Bernardino  county,  California,  occurs  a  copper 
bearing  gold  ore  containing  the  copper  in  the  form  of  silicate.  The 
silicate  of  copper,  being  soluble  in  cyanide,  causes  a  loss  of  7  to  8  Ib. 
cyanide  per  ton  of  ore  by  the  ordinary  treatment.  By  Hunt 's  method 
8  Ib.  quick-lime  is  added  per  ton  of  ore.  The  ore  in  the  vat  is  then 
leached  with  a  0.15%  solution  of  cyanide  to  which  has  been  added 
6  Ib.  ammonium  chloride  per  ton  of  solution.  The  solution  after 


OF    THE    COMMON    METALS.  209 

contact  with  the  ore  is  drained,  and  the  ore  given  six  days' 
continuous  contact  and  washing.  The  gold-bearing  solution  can 
be  precipitated  in  zinc-boxes,  but  electrolytic  precipitation  is 
preferred.  In  the  precipitation,  lead  anodes  and  aluminum  cathodes 
are  desired.  The  anodes  are  peroxidized  by  dipping  into  a  solution 
of  potassium  permanganate  before  use.  A  current  density  of  3 
amperes  per  square  foot  is  employed.  The  gold,  silver,  and  copper 
do  not  precipitate  as  an  adhering  coating  upon  the  aluminum 
cathodes,  but  fall  from  them  as  sludge  to  the  bottom  of  the  tank. 
The  method  has  been  successfully  applied  to  the  treatment  of 
copper-bearing  mill-tailing  dumps  exposed  for  years  to  the  weather. 
Potassium  or  sodium  cyanide  costs  18  to  20c.  per  pound.  Ammonium 
chloride,  commercially  called  muriate  of  ammonia,  costs  as  an 
estimate  5  to  6c.  per  pound. 

48.     CYANIDATION  OF  CONCENTRATE. 

Two  cases  arise  in  respect  to  the  treatment  of  the  sulphide 
concentrate  obtained  from  gold  or  silver  ores.  In  the  first,  the 
valuable  metals  enter  the  concentrate,  while  the  gangue  is 
comparatively  barren.  In  this  case  only  a  small  percentage  of  the 
total  ore  need  be  cyanided.  In  the  second,  all  parts  of  the  ore 
have  value.  Here  we  have  the  choice  between  finely  grinding  all 
the  ore  for  extraction  or  on  the  other  hand  treating  a  large  part 
of  the  ore  by  ordinary  methods  of  extraction,  but  giving  the 
concentrate,  which  is  of  higher  grade,  and  in  smaller  quantity,  a 
more  careful  treatment.  The  removal  of  the  concentrate  from  the 
ore  frees  the  tailing  from  acid  sulphates  and  base  metals  which 
would  interfere  with  cyanidation. 

Gold  and  silver-bearing  concentrate  can  be  cyanided  with  a 
high  extraction  in  some  cases.  The  gold  must  be  fine  and  the  silver 
present  as  silver  chloride  or  sulphide.  When  silver  is  the  principal 
metal  to  be  extracted,  the  treatment  must  be  prolonged.  Thus  with 
ore  re-ground  to  100-mesh  size,  an  extraction  of  93.5%  of  the  silver 
and  96%  of  the  gold  was  obtained  in  one  instance  on  ore  of  130  oz. 
Ag  and  1  oz.  Au  per  ton.  This  was  accomplished  by  an  eight-day 
treatment  by  agitation  and  decantation,  which  is  equivalent  to 
leaching  30  days.  Ror  high-grade  impure  concentrate  the  best 
method  is  to  give  a  chloridizing  roast,  following  with  a  careful  and 
prolonged  leaching  with  cyanide  solution.  The  silver  compounds 
are  thus  converted  into  a  readily-soluble  form,  and  the  ore  is 
rendered  porous  and  penetrable  by  the  solution. 

Treatment    at    the    Standard    plant,    Bodie,    California. 


210  THE    METALLURGY 

concentrate,  consisting  of  iron  oxide  with  a  little  pyrite  is  charged 
in  one-ton  lots  into  an  ordinary  5-ft.  pan  (see  Fig.  103),  lime  and 
water  first  having  been  added  to  dilute  the  pulp  to  45%  solid  matter. 
The  charge  is  ground  48  hours,  then  cyanide  is  added  in  quantity 
sufficient  to  strengthen  the  solution  to  1.2%.  Grinding  is  continued 
24  hours  with  additions  at  intervals  of  quick-lime  and  cyanide.  The 
lime  insures  alkalinity,  the  cyanide  maintains  the  strength  of  the 
solution  at  a  fixed  percentage.  During  the  grinding  an  oxidizing 
action  occurs  and  the  pulp  changes  in  color  from  a  greenish  shade 
to  a  brownish  red.  It  is  drawn  into  a  settler,  maintained  in  motion 
for  24  hours,  and  is  diluted  with  weak  cyanide  solution,  filling  the 
settler.  The  whole  content  is  then  run  to  the  filter  press.  The 
filtrate  from  the  press,  containing  the  gold  in  solution,  is  precipi- 
tated with  zinc-shaving. 

The  extraction  of  metals,  96.8%  of  the  gold  and  84.1%  of  the 
silver,  is  high.  The  cost  also,  $7.91  per  ton,  is  high,  and  18  to  20 
Ib.  cyanide  is  consumed  per  ton  of  concentrate  treated. 

Treatment  at  Harrisburg,  Arizona. — Seventy  tons  of  ore  daily 
were  concentrated,  yielding  two  tons  of  concentrate,  while  the  mill- 
tailing  being  of  little  value  was  thrown  away.  The  raw  concentrate 
is  exposed  to  the  air  for  a  while  to  partly  oxidize  it,  since  it  is 
found  that  when  thus  treated  it  grinds  more  easily,  and  the 
extraction  is  better.  It  is  then  charged  to  a  5-ft.  pan,  the  charge 
consisting  of  1.5  tons  of  concentrate  and  1  ton  of  solution  containing 
6  Ib.  quick-lime  and  6  Ib.  potassium  cyanide.  The  grinding,  which 
takes  8  hp.,  was  continued  8  hours  after  adding  2  Ib.  more  cyanide, 
the  temperature  is  increased  some  40°.  The  content  of  the  pan, 
not  too  finely  ground,  is  now  run  into  a  15-ton  leaching  tank. 
When  the  slime  is  partly  settled  in  the  vat,  dry  tailing  is  sprinkled 
in.  This  coarse  material  facilitates  percolation,  and  we  thus  have  a 
number  of  layers  of  charge  interstratified  with  middling.  The 
tank  having  been  filled,  the  charge  is  continuously  leached  with 
cyanide  solution  while  the  other  tank  is  being  filled.  Thus  each 
tank  is  7  days  filling,  followed  by  4  days  leaching,  and  3  days 
washing,  before  it  is  discharged.  The  average  extraction  is  94%, 
and  the  consumption  of  cyanide  is  8  Ib.  per  ton.  The  cost  of 
treatment  of  the  concentrate  is  estimated  $5  per  ton. 

49.     THE  BROMO-CYANOGEN  PROCESS. 

This  method,  called  also  the  'Diehl  process,'  is  practised  upon 
the  sulpho-telluride  ores  of  Western  Australia.  A  typical  ore 
contains  50%  Si02 ;  10%  Fe;  7  to  22%  CaO  and  MgO,  3  to  7%  S; 


OF    THE    COMMON    METALS.  211 

0.03  to  0.10%  Te,  and  2  to  3  oz.  gold  per  ton.  The  ore  contains 
iron,  calcium,  and  magnesium  carbonates  which  give  it  cement-like 
properties  after  roasting,  disposing  it  to  cake  or  set  when  subjected 
to  the  action  of  cyanide  solution  in  the  tank.  Since  the  Diehl 
process  is  applied  to  the  raw  ore,  this  trouble  is  avoided,  and  the 
process  has  became  a  successful  method  of  treating  a  certain  kind 
of  ore. 

Fine  grinding  and  amalgamation  was  tried  upon  the  ore,  but 
an  extraction  of  only  20%  resulted.  Upon  roasting  and  amal- 
gamating in  pans,  the  extraction  was  44%.  Concentration  gave 
poor  results  owing  to  the  loss  of  the  finely  ground  almost  inpalpable 
tellurides  escaping  with  the  tailings.  Ordinary  cyanidation  gave 
extractions  of  60  to  77%,  but  with  preliminary  roasting,  it  yielded 
a  total  of  93  to  94  per  cent. 

It  is  found  that  the  ore  must  be  finely  ground,  or  slimed  as 
it  may  be  termed,  to  obtain  a  good  extraction.  At  the  same  time 
we  must  distinguish  between  a  product  that  contains  many  fine  hard 
particles  of  quartz  and  one  consisting  of  real  slime  composed  of 
colloid  substances,  in  which  no  grit  exists.  Moreover,  the  above  tests 
show  that  only  a  part  of  the  gold  of  the  ore  is  in  a  metallic  state, 
the  rest  being  present  as  telluride  and  locked  within  the  crystals  of 
pyrite.  Owing  to  the  brittleness  of  the  telluride,  the  finer  the  slime 
the  richer  is  the  content  of  gold. 

The  process. — The  ore  is  coarsely  crushed  and  fed  to  the  stamps. 
The  pulp  from  the  stamps  is  classified,  the  overflow  going  to  the 
agitation  vats,  while  the  underflow,  of  coarse  sand,  goes  to  the 
tube-mills  for  re-grinding.  In  the  tube-mills  the  sand  is  crushed 
so  that  most  of  it  passes  a  200-mesh  screen.  The  product,  still 
containing  5%  sandy  particles,  is  again  classified,  the  overflow  as 
before  going  to  the  agitation  vats,  while  the  sand  is  returned  to 
the  tube-mill  for  re-crushing.  Thus  the  agitation  vats  recover  only 
a  slimed  product,  fine  enough  to  pass  a  200-mesh  screen.  The 
agitation  vat  (See  Fig.  101)  is  25  ft.  diam.  by  8  ft.  high.  When 
it  has  been  filled,  potassium  cyanide  is  added  to  produce  a  0.22% 
KCN  solution.  This  is  used  in  the  proportion  of  two  of  solution 
to  one  of  the  dry  slime,  and  the  whole  is  stirred  for  a  period  varying 
between  16  and  24  hours.  This  is  followed  by  the  addition  of 
bromo-cyanogen  to  increase  the  strength  of  the  solution  by  0.055% 
salt  used  per  ton  of  dry  slime.  Agitation  is  continued  until  the 
gold  is  in  solution,  after  which  the  whole  is  filter-pressed.  Since, 
by  this  method  of  final  treatment,  but  little  water  is  required,  the 


212 


THE    METALLURGY 


process  has  an  important  advantage  in  the  country  mentioned  where 
water  is  worth  50c.  per  1000  gallons. 

The  pulp  is  pumped  to  presses  of  the  Dehne  type  that  have  a 
capacity  of  three  to  five  tons  of  dry  slime  per  charge,  the  thickness 
of  the  cake  in  the  recesses  varying  from  one  to  three  inches.  The 
distance-frames  in  which  the  cakes  form  are  40  in.  square,  there 
being  30  to  50  cakes  in  a  press.  Of  the  solution  70%  is  recovered 
at  once ;  then  a  weak  solution  or  water  is  forced  through,  and 
complete  washing  accomplished  by  the  use  of  a  half  ton  of  solution 
per  ton  of  slime.  Compressed  air,  under  90  Ib.  pressure  is  next 


Fig.    101.      AGITATING  VAT. 

introduced.  This  soon  expels  the  water  from  the  cakes  leaving 
only  11  to  15%  moisture.  The  press  being  opened,  the  cakes  are 
discharged  into  cars  below  and  trammed  to  the  dump.  It  takes 
two  men  30  to  45  minutes  to  discharge  a  press,  clean  the  frames, 
and  close  it,  ready  for  another  charge.  The  time  of  filling,  using 
a  pressure  of  60  to  100  Ib.  per  sq.  in.,  is  20  to  30  minutes.  Wash- 
water  is  used  15  to  25  minutes,  and  compressed  air  for  10  minutes. 
The  full  cycle  of  operations  can  be  completed  therefore  in  two 
hours.  Two  men,  using  two  presses,  handle  seven  charges  per 
8-hour  shift,  or  40  tons  dry  slime  per  press  in  24  hours.  The 
extraction  by  the  use  of  the  press  is  90  per  cent. 

Bromo-cyanide   is  made  by  dissolving  the  commercial  crystals 
in  water,  or,  when  produced  at  the  works,  by  adding  bromine-water 


OF    THE    COMMON    METALS.  213 

to  cyanide  solution.  The  cyanide  reacts  with  potassium  bromide 
as  follows : 

KCN  +  BrCN  =  KBr  +  C2N2 

Cyanogen  (C2N2)  in  nascent  condition  acts  energetically  on  the 
gold  as  follows : 

2Au  +  C2N2  +  2KCN  =  2KAu(CN)2 

Bromo-cyanide  does  not  require  the  prior  aeration  necessary  in 
the  use  of  the  ordinary  solution.  It  is  expensive,  however,  and 
has  been  applied  only  to  sulpho-tellurides  that  are  unattacked, 
or  are  attacked  but  slowly  by  plain  cyanide. 

50.     SMELTING  OF  GOLD  ORES. 

Gold  may  be  recovered  from  its  ore  by  the  processes  of  silver- 
lead,  or  of  copper-matte  smelting.  It  often  is  found  in  copper  or 
lead-bearing  ores  and  when  in  excess  of  0.05  oz.  per  ton,  is  paid  for 
by  the  smelting  works  at  the  rate  of  $19  to  $19.50  per  ounce. 
Practically  all  the  gold  is  recovered  in  smelting,  and  this  would  be 
the  best  method  of  treatment  were  it  not  for  the  high  cost  of 
freight  and  of  treatment.  If  smelted  near  the  mine  in  a  works 
operated  by  the  mining  company,  the  cost  of  freight  is  eliminated. 
The  charge  for  dry,  fairly  silicious  ores,  from  Cripple  Creek  and 
from  Boulder  county,  Colorado,  is  from  $4  to  $10  per  ton,  according 
to  grade.  The  low-grade  ores  are  subject  to  a  low  treatment  rate. 
On  the  other  hand  ore  treated  by  milling  and  amalgamation,  or  by 
cyanidation,  while  the  extraction  is  less,  often  yields  higher  net 
returns.  An  example  is  found  in  the  case  of  the  dry  silicious  gold 
ore  from  Boulder  count}7,  Colorado,  containing  0.5  oz.  Ag  per  ton, 
giving  70%  extraction  by  milling  and  amalgamation,  or  of  90% 
by  cyanidation.  In  comparing  the  costs  we  have : 

SMELTING. 

100%  of  0.5  oz.  Au  at $19.00              $9.50 

Mining     2.00 

Freight    1.50 

Treatment  4.00                 7.50 


Net  returns  $2.00 

MILLING  AND  AMALGAMATION. 

70%    of  0.5   oz.   Au   at $20.50  $7.17 

Mining 2.00 

Milling    1.00  3.00 

Net   returns  $4.17 


214  THE    METALLURGY 

CYANIDATION. 

90%   of  0.5  oz.  Au  at $20.50  $9.23 

Mining     2.00 

Cyaniding     1.65  3.65 


Net  returns     $5.58 

From  the  above  comparison  it  is  seen  that  cyaniding  is  the 
most  profitable  method  of  treatment  for  this  grade  of  ore,  and  at 
this  place. 


PART  IV.     SILVER 


/*• 


PART  IV.  SILVER. 

51.     SILVER   ORE. 

The  silver  minerals  of  importance  in  treatment  are  as  follows : 

Native  silver  sometimes  occurs  in  the  form  of  flakes  or  leaves, 
and  as  wire-silver  and  metallic  silver  adherent  to  native  copper. 
Native  silver  can  be  readily  amalgamated,  but  when  present  in 
particles  of  visible  size  it  is  so  slowly  soluble  in  cyanide,  that 
practically  no  extraction  can  be  obtained. 

Cerargyrite  (horn-silver,  silver  chloride),  AgCl,  is  widely 
distributed.  At  mines  it  is  found  in  the  upper  oxidized  zones. 
It  is  probable  that  much  of  the  so-called  chloride  ore  is  really 
a  chloro-bromide  (embolite).  The  ore  is  readily  amalgamated  and 
is  free-milling.  The  silver  also  is  readily  soluble  in  cyanide  and 
in  sodium  hyposulphite  solutions. 

Argentite,  Ag2S,  is  one  of  the  common  silver  ores.  By  using 
chemicals  (bluestone  and  salt)  it  can  be  amalgamated  in  pans,  and 
the  silver  extracted  thus  from  the  ore.  It  is  soluble  in  potassium 
cyanide  solution. 

Stephanite,  5Ag2S,Sb2S3 ;  pyrargyrite,  3Ag2S,Sb2S3 ;  proustite, 
3Ag2S,As2S3 ;  dycroasite,  Ag3Sb  are  silver  sulph-arsenides  or 
sulph-antimonides,  refractory  in  amalgamation,  even  with  chemicals, 
sparingly  soluble  in  cyanide  solution,  but  readily  soluble  in  a 
solution  of  mercurous  potassic  cyanide. 

Finally  we  have  those  silver  sulphides  that  contain  also  copper. 
These  are  polybasite,  9(Ag2Cu)S(SbAs)2S3,  and  tetrahedrite 
(gray  copper  ore,  fahlerz),  4CuFeAg2(HgZn)S,(SbAs)S3  the 
most  complex  of  all,  in  which  the  silver  varies  from  0.06  to 
31%,  being  higher  in  the  arsenical  and  lower  in  the  antimonial 
varieties.  These  sulphides  are  refractory  to  any  amalgamation 
method,  and  because  of  their  copper  content  are  precluded  from 
treatment  by  cyanide,  even  when  roasted.  This  does  not  interfere 
with  treatment  by  hyposulphite  lixiviation  after  roasting. 

A  number  of  rare  minerals  containing  silver  could  also  be 
enumerated,  but  for  the  metallurgist  the  minerals  above  named  are 
the  important  ones. 


218  THE    METALLURGY 

Silver  ores  in  general  contain  but  a  small  percentage  of  precious 
metal.  They  are  composed  mostly  of  gangue  (waste  matter  of  the 
ore),  and  many  are  treated  that  contain  less  than  0.1  to  0.2%  silver. 
Thus  we  have  at  the  Comstock  Lode,  Nevada,  silver  in  native  form 
and  as  sulphide,  but  oxides  of  iron  and  manganese  with  the 
associated  sulphides,  pyrite,  blende,  galena,  and  chalcopyrite.  At 
the  Ontario  mine,  Park  City,  Utah,  the  silver  occurs  as  argentite 
and  tetrahedrite  in  a  gangue  of  quartz  and  clay  associated  with  a 
little  of  the  heavy  minerals  blende  and  galena.  These  sulphides 
carry  silver  which  is  recovered  with  the  concentrate  in  case  of 
concentration. 

52.     THE   EXTRACTION    OF    SILVER   FROM    ORES. 

Silver  is  extracted  from  its  ores  by  amalgamation  in  pans,  by 
hydro-metallurgical  methods,  and  by  smelting.  Amalgamation  is 
now  used  less  in  America  than  formerly,  smelting  having  taken 
the  place,  not  only  because  of  better  extraction  and  lower  treatment 
rates,  but  because  ores  are  more  easily  shipped  to  smelting  works 
in  result  of  the  extension  of  railroad  facilities.  Certain  ores 
contain  the  silver  in  a  form  suitable  for  cyaniding,  and  the  hydro- 
metallurgical  method  in  consequence  is  coming  forward.  The  other 
methods  have  partly  dropped  out  of  use.  No  reason  appears  why 
hyposulphite  lixiviation  should  not  revive  under  the  stimulus  of 
the  recent  methods  of  agitation  and  filter-pressing.  The  patio 
process,  formerly  much  practised  in  Mexico  where  conditions 
favored,  has  been  superceded  by  cyaniding  in  many  cases,  on 
account  of  the  lower  cost  of  operating  the  latter  process,  but  in 
the  past,  large  quantities  of  silver  have  been  extracted  by  the  patio 
process. 

53.     SILVER  MILLING  AND  AMALGAMATION. 

The  silver  ores  suitable  to  treat  by  milling  and  amalgamation 
are  those  that  contain  the  metal  in  such  form  as  to  be  acted  upon 
by  mercury  when  assisted  by  agitation,  heat,  and  certain  chemicals. 
The  ore  is  first  crushed  fine  by  stamps,  as  in  gold  milling,  then 
treated  for  several  hours  in  pans,  the  reactions  being  slow  compared 
with  those  of  the  amalgamation  of  gold.  In  gold  milling,  the 
greater  part  of  the  gold  can  be  arrested  on  an  apron-plate  during 
the  few  seconds  in  which  the  ore  is  passing;  while  in  silver  milling, 
the  ore-pulp  has  several  hours  contact  with  mercury  aided  by  heat 
and  chemicals,  and  is  but  slowly  amalgamated.  In  gold  milling, 
ore  containing  0.5  oz.  Au  per  ton  can  be  profitably  milled.  In 


OF    THE    COMMON    METALS.  219 

silver  milling,  ore  of  equivalent  value  would  contain  20  oz.  Ag  per 
ton,  or  40  times  as  much  metal.  Thus  is  seen  why  so  much  time 
is  allowed  in  silver  milling,  and  why  so  many  precautions  must 
be  taken  to  be  sure  that  all  metal  possible  is  recovered.  Several 
ounces  of  silver  per  ton  often  remain  in  the  tailing. 

The  silver  metals  suited  to  pan  amalgamation  are  cerargyrite 
(horn  silver,  silver  chloride),  native  silver  in  flakes,  wire,  or  other 
forms,  and  certain  silver  sulphides,  notably  argentite  (Ag2S).  When 
the  ore  is  refractory,  containing  arsenical  and  antimonial  sulphides, 
and  especially  containing  tetrahedrite,  galena,  or  blende,  it  is 
necessary  to  roast  with  salt,  setting  free  the  silver  or  converting 
it  into  the  form  of  a  chloride,  which  becomes  susceptible  to 
amalgamation.  There  is  no  sharp  line  of  demarkation  between  free- 
milling  and  roasting  milling  ores.  Often  the  upper  part  of  a  vein  is 
free-milling.  In  depth  base  metals  and  sulphides  begin  to  come  in, 
and  it  finally  becomes  necessary  to  roast  the  ore.  The  best 
extraction  therefore  is  obtained  from  decomposed  or  oxidized  ore, 
in  which  the  silver  minerals  occur  in  a  form  that  renders  possible 
the  action  of  the  mercury.  There  are  few  deposits  of  oxidized  ores 
containing  silver  chloride  and  native  silver  that  as  a  whole  are 
suitable  for  free  silver  milling.  Such  ore,  so  far  as  silver  chloride 
is  concerned,  can  also  be  treated  by  cyanidation,  but  the  latter 
method  would  not  recover  native  silver. 

Arsenic  and  antimony  compounds  interfere  with  amalgamation 
by  fouling  the  quicksilver,  checking  the  reactions  of  the  chemicals 
added  to  promote  amalgamation,  and  by  carrying  off  silver,  which 
is  incapable  of  being  amalgamated  with  them. 

We  may  divide  the  methods  of  silver-milling  into 

(1)  Wet  silver-milling,  or  the  Washoe  process,  which  includes 
(a),  the  tank  mill,  and  (b)  the  Boss  process. 

(2)  Combination    process,    combining    the    gold-mill    and    the 
silver-mill. 

(3)  Dry  silver-milling,  or  the  Reese  River  process. 

54.     WET  SILVER-MILLING  WITH  TANK  SETTLING. 

This  is  also  known  as  the  Washoe  process,  receiving  the  name 
from  the  place  where  it  was  perfected  for  the  treatment  of  ores 
from  the  Comstock  Lode,  Nevada.  The  process  is  applicable  to 
the  so-called  free-milling  ores,  in  which  the  silver  occurs  native, 
as  chloride  or  in  small  amount  as  argentite.  The  ore  should  be 
free  from  lead  and  from  any  tough  clayey  gangue. 

In  wet  silver-milling,  when  settling  tanks  are  used,  the  process 


OF    THE    COMMON    METALS.  221 

consists  in  coarse-crushing  the  ore,  stamping  it  fine,  and  collecting 
it  in  settling  tanks.  The  crushed  sand  is  ground  in  amalgamating 
pans  using  mercury  to  collect  the  silver.  The  sand  is  separated 
from  the  silver-bearing  mercury  in  settling  pans  and  is  rejected. 
The  amalgam  is  strained  from  the  mercury,  retorted,  and  the  retort- 
residue  melted  into  silver  ingots.  Gold  present  in  the  ore  is 
recovered  as  well  as  the  silver.  The  process  resembles  gold  milling 
except  that  amalgamation  and  the  removal  of  the  amalgam  is 
effected  in  pans. 

Fig.  102  is  a  sectional  elevation  of  a  wet-crushing  tank-mill 
for  the  treatment  of  free-milling  silver  ores.  The  ore  from  the 
mine,  as  in  gold  milling,  is  dumped  from  the  tram-car  over  a  grizzly 
or  bar-screen  a,  the  bars  being  set  1*4  in.  apart.  The  fine  falls 
into  the  sloping  bottom  storage  bin  c,  while  the  lump  ore  is  received 
upon  the  feed-floor  at  the  mouth  of  the  Blake  crusher  6,  and  after 
being  coarsely  crushed,  joins  the  fine  in  the  bin.  The  coarse 
crushing  is  done  during  the  10-hour  day-shift. 

The  ore  passing  to  the  battery  is  fed  by  an  automatic  feeder 
d  such  as  is  used  in  gold  milling.  The  flow  of  ore  to  the  automatic 
feeder  is  controlled  by  gates  at  the  inclined  outlet  chutes  and  the 
feeder  is  kept  full  to  supply  the  stamps  e  uniformly  as  required. 
Water  (6  to  8  tons  per  ton  of  ore)  is  at  the  same  time  supplied  in 
the  mortar,  and  the  suspended  ore  forms  a  pulp,  which  is  splashed 
through  the  30-mesh  screens  by  the  motion  of  the  stamps.  A  double- 
discharge  mortar  of  the  type  suited  to  silver  milling  is  seen  at 
Fig.  48.  The  large  screen-opening  possible  with  a  double  discharge 
favors  a  more  rapid  pulverization  than  would  be  possible  with  a 
single-discharge  mortar.  (For  further  particulars  of  gravity  stamps 
see  the  detailed  description  under  'Gold  Milling.')  The  pulp  flows 
by  launders  /  into  the  settling  boxes  or  tank  g  1  ft.  square  by  3  ft. 
deep.  There  is  a  double  row  of  these  tanks,  20  in  a  row,  occupying 
the  length  of  the  mill  in  front  of  the  stamps.  The  flow  of  the 
pulp  is  from  box  to  box  until  it  goes  by  launder  to  a  settling-pond 
outside  the  mill.  Most  of  the  solids  settle  in  the  first  boxes,  a 
further  portion  dropping  in  the  succeeding  ones,  and  the  turbid 
water  passing  to  the  pond.  Here  it  has  its  final  chance  to  settle 
before  running  to  waste,  or  it  may  be  again  used  in  the  mill  if 
water  is  so  scarce  that  it  pays  to  do  this.  The  settled  slime  is 
dug  from  the  pond  at  a  later  time  and  treated  like  the  rest  of  the 
crushed  ore. 

A  variation  of  this  method,  shown  in  Fig.  102,  consists  in 
conducting  the  flow  from  the  last  box  g'  by  an  inclined  elevator  to 


222 


THE    METALLURGY 


a  tank  h  situated  in  front  of  and  above  the  battery,  the  dirty  water 
being  again  used  for  stamping.  When  the  first  settling  box  is 
full  the  flow  of  pulp  is  by-passed  into  the  next  one.  The  contents 
of  the  full  box  is  shoveled  upon  the  floor  adjoining,  and  thence 
taken  as  needed  to  the  amalgamating-pans  q.  The  emptied  box 
has  the  flow  of  the  last  one  turned  into  it,  thus  making  it  the 
last  in  the  series,  and  the  launders  are  so  arranged  that  this  can 
be  done. 

The  ore,  thrown  out  upon  the  floor,  is  fed  directly  into  the  pan, 


Fig.    103.      COMBINATION  AMALGAMATING  PAN. 

or   loaded   into   the   tram-car   seen   in   Fig.    102,   and   conveyed   to 
the  pan. 

Fig.  103  represents  two  of  these  pans.  One  is  shown  in  section, 
the  other  in  elevation.  The  pan  shown  in  section  is  5  ft.  diam.  by 
30  in.  deep,  and  has  a  cast-iron  bottom  and  the  sides  made  of 
wooden  staves.  It  is  furnished  with  a  central  sleeve  or  cone  through 
which  rises  a  shaft  carrying  a  cylindrical  casting  called  a  spider, 
which  becomes  bell-shaped  and  broadens  into  feet  below.  The 
spider  carries,  bolted  to  the  feet,  a  flat  cast-iron  ring  called  a 
muller,  and  to  the  underside  of  the  muller  is  attached  six  shoes 


OF    THE    COMMON    METALS.  223 

or  plates  of  chilled  cast-iron  21/2  in-  thick.  The  spider,  muller,  and 
shoes  are  raised  or  lowered  as  desired,  by  means  of  a  hand-wheel  and 
screw  at  the  top  of  the  shaft,  which  is  driven  by  bevel  gearing 
from  the  horizontal  shaft  and  pulley  below.  Upon  the  bottom  of 
the  pan  rest  chilled  cast-iron  plates  or  dies  that  furnish  the  lower 
or  fixed  grinding  surface.  The  shoes  attached  to  the  muller  revolve 
60  rev.  per  min.,  and  rubbing  upon  the  dies,  grind  the  ore. 

In  working  the  pans,  the  shoes  are  raised  %  in-  from  the  dies 
and  set  in  motion,  the  pan  is  partly  filled  with  water,  and  3000 
Ib.  of  the  damp  pulverized  ore  are  shoveled  in.  The  ore  and 
water  nearly  fill  the  pan  and  the  mixture  is  stirred  until  it  is 
of  the  consistence  of  honey.  The  motion  establishes  a  movement 
or  current  of  pulp  beneath  the  muller  toward  the  periphery.  At 
the  periphery  it  rises,  flows  toward  the  center,  sinks,  and  passes 
again  under  the  shoes.  To  assist  the  action,  the  rising  pulp  is 
deflected  inward  by  cast-iron  wing-plates. 

After  thorough  mixing  in  the  pan  the  shoes  are  lowered  until 
they  touch  the  dies,  and  grinding  goes  on  for  1%  hours,  the  content 
of  the  pan  being  meanwhile  heated  nearly  to  boiling  by  steam  under 
pressure  from  a  pipe  that  dips  beneath  the  surface  of  the  charge, 
the  pan  being  covered.  Sometimes  the  pan  is  provided  with  a  double 
bottom  into  which  exhaust  steam  from  the  engine  is  introduced. 

After  grinding,  the  shoes  are  raised  and  300  Ib.  mercury  (10% 
the  weight  of  the  ore)  is  added,  by  sprinkling  it  through  a  fine 
strainer.  The  mixing  is  then  continued  four  hours.  The  mercury 
takes  up  silver  most  rapidly  at  first,  but  the  action  afterward 
slackens.  At  the  end  of  the  first  hour,  in  one  instance,  74.7% 
was  amalgamated,  of  the  second  hour  76.3%,  of  the  third  77.7%, 
and  at  the  end  of  the  fourth,  81.0%  the  silver  contained  in  the 
ore.  After  the  fourth  hour  no  further  silver  amalgamated.  The 
globules  of  mercury  suspended  in  the  pulp  take  the  silver  as  they 
come  in  contact  with  it.  Care  is  taken  to  have  the  pulp  of  the 
right  consistence  so  that  mercury  will  not  settle  out.  This  condition 
is  shown  when  a  wooden  stick,  dipped  in  the  pulp  and  withdrawn, 
is  found  to  be  covered  with  a  thick  mud  in  which  are  disseminated 
minute  globules  of  mercury.  If  the  ore  is  refractory,  salt  and 
copper-sulphS^  are  advantageously  added  at  the  beginning  of 
grinding  to  accelerate  the  reactions,  promote  amalgamation,  and 
increase  the  yield  of  silver. 

The  time  required  for  grinding  varies  from  four  to  six  hours. 
In  mills  treating  chloride  ore,  the  charges  are  finished  in  four 
hours  without  grinding.  In  treating  refractory  ores,  the  total 


224  THE    METALLURGY 

time  for  a  charge  becomes  six  to  eight  hours,  of  which  time  four 
hours  are  required  for  grinding. 

The  charge  above  treated  having  been  amalgamated,  the  pan  is 
ready  to  empty  into  the  settler  r.  About  15  minutes  before  the 
discharging,  the  speed  of  the  muller  is  reduced  to  40  rev. 
per  min.  and  the  pan  filled  to  the  top  with  water.  A  plug 
closing  the  discharge  opening  at  the  bottom  of  the  pan,  seen  at 
the  left  in  section,  Fig.  103,  is  pulled  out,  and  the  entire  content 
run  by  launder  to  an  8-ft.  settler,  at  a  lower  level,  shown  at  the 
right  of  the  amalgamating  pans  in  Fig.  102.  Emptying  the  pan 
and  washing  it  with  a  hose  takes  half  an  hour,  after  which  time 
the  plug  is  replaced,  and  the  pan  is  ready  for  another  charge. 
Thus  the  total  time  for  the  cycle  of  operations  described  is  six 
hours,  making  it  possible  to  treat  four  charges  daily. 

The  reactions  that  take  place  in  the  pan  are  as  follows  : 

Native  silver  in  threads,  films,  flakes,  or  grains  readily  combines 
with  the  mercury  and  forms  an  amalgam  which  contains  a  large 
excess  of  mercury. 

Silver  chloride  in  contact  with  the  mercury  decomposes  as 
follows : 

(1)  2AgCl  +  2Hg  =  Hg2Cl2  +  2Ag 

The  metallic  silver  liberated  amalgamates  with  additional 
mercury.  The  particles  of  iron,  abraided  from  the  stamps  and  the 
bottom  of  the  pan,  decompose  the  mercury  salt  and  liberate  the 
mercury  as  follows : 

(2)  Hg2Cl2  +  Fe  =  FeCl2  +  2Hg 

In  many  so-called  free-milling  silver  ores  argentite  is  contained 
which  in  part  is  decomposed  by  mercury  as  follows : 

(3)  Ag2S  +  2Hg  =  Ag2  +  HgS 

The  sulphide  of  mercury  thus  formed  is  lost.  We  have  already 
stated  that  chemicals,  notably  copper  sulphate  and  common  salt, 
are  added  to  promote  the  decomposition  of  the  silver  sulphide. 
There  is  added  in  the  amalgamating  pan  from  6  to  18  Ib.  salt  and 
from  3  to  9  Ib.  copper  sulphate  per  ton  of  ore  treated.  The  reactions 
as  generally  given  are  the  following: 

(4)  CuS04  +  2NaCl  =  Na2S04  +  CuCl2 

The  chloride  of  copper  acting  on  the  silver  sulphide  decomposes  it : 

(5)  Ag2S  +  CuCl2  =  CuS  +  2AgCl 

The  silver  chloride  amalgamates  as  shown  by  reaction   (1). 


OF    THE    COMMON    METALS. 


225 


The  complete  separation  of  the  mercury  with  the  silver-amalgam 
is  effected  in  the  settler,  there  being  one  settler  provided  for  two 
amalgamating  pans.  The  settler  is  a  cylindrical  pan  8  ft.  diam.  by 
3  ft.  deep,  3  times  the  capacity  of  the  amalgamating  pan,  but  of 
similar  construction,  as  shown  in  Fig.  104.  It  is  required  to  do  no 
grinding,  but  to  gently  agitate  the  pulp  with  the  wooden  shoes 
with  which  it  is  provided.  The  shoes  nearly  touch  the  bottom  of 


Fig.    104.      EIGHT-FOOT   SETTLER. 

the  settler,  the  exact  height  being  adjustable.  The  grooved  border 
at  the  bottom  just  within  the  sides  of  the  settler  has  a  slight  grade 
to  the  outlet  and  mercury-well  at  the  left.  The  mercury  settles 
from  the  pulp,  flows  to  the  lowest  point  and  stands  at  a  height 
that  balances  the  hydrostatic  head  of  the  content  of  the  pan.  Since 
the  specific  gravity  of  mercury  is  14  and  the  content  of  the  settler 
approximately  1.5  the  height  of  the  mercury  is  a  little  less  than 
4  inches.  The  bottom  outlet-hole  of  the  well  is  plugged.  At 


226  THE    METALLURGY 

different  heights  in  the  side  of  the  pan  there  are  provided  openings 
that  are  kept  closed  by  plugs.  When  the  plugs  are  withdrawn 
the  tailing  and  water,  free  from  mercury,  passes  out  of  the  pan. 

The  shoes  of  the  settler  having  been  set  in  motion,  at  the  rate 
of  15  rev.  per  min.,  and  raised  8  in.  above  the  bottom,  the  contents 
of  the  two  pans  are  run  in,  as  has  been  described.  Water  is  then 
added  to  within  6  in.  of  the  top,  greatly  thinning  the  pulp,  and 
filling  the  settler.  After  half  an  hour  the  shoes  are  gradually 
lowered  until,  at  the  end  of  two  hours,  they  nearly  touch  the 
bottom.  The  purpose  of  the  agitation  is  to  keep  the  lighter  portion 
of  the  ore  (now  called  the  tailing)  in  suspension,  while  the  silver- 
bearing  mercury,  the  heavier  particles  of  sulphide,  and  the  particles 
of  iron  from  the  stamps  collect  at  the  bottom.  The  stirring  is 
continued  3%  hours,  after  which  the  highest  plug  in  the  side  of 
the  settler  is  removed,  and  the  turbid  water  containing  tailing  is 
allowed  to  escape  by  launder,  a  stream  of  clear  water  being  mean- 
while allowed  to  flow  through.  The  plugs  are  then  withdrawn  one 
by  one  until  the  settler  is  emptied  of  all  the  content  except  the 
heavy  portion  containing  sulphide,  iron  particles,  and  the  mercury. 
Emptying  takes  half  an  hour,  and  the  cycle  of  operations  becomes 
six  hours  as  in  the  case  of  the  amalgamating  pan.  Since  escaping 
tailing  contains  sulphide,  it  may  be  run  over  riffles,  or  blanket- 
lined  launders,  before  running  to  waste.  In  Fig.  109  is  see  a 
system  of  pans  and  settlers. 

The  silver-bearing  mercury  or  diluted  amalgam,  a  mixture  of 
silver-amalgam  and  mercury,  collecting  in  the  mercury  well,  over- 
flows by  an  escape-opening  indicated  in  Fig.  104.  From  the  opening 
it  passes  by  a  half-inch  pipe  to  the  amalgam  safe  shown  at  the 
right  of  the  settler,  Fig.  102.  The  safe,  arranged  to  prevent 
theft  of  the  amalgam,  is  shown  on  a  larger  scale  in  Fig.  105.  The 
amalgam  and  mercury  enter  a  conical  canvas  sack  or  filter.  The 
mercury  oozes  through  the  pores  of  the  canvas  while  the  amalgam 
containing  as  little  as  14%  silver  is  retained.  Occasionally,  after 
amalgam  has  accumulated,  the  sack  is  squeezed  between  the  hands 
to  remove  the  surplus  mercury,  and  the  compressed  amalgam 
containing  20  to  28%  silver,  is  reserved  for  retorting.  The  mercury 
flows  out  at  the  bottom  through  an  outlet  provided,  as  seen  in 
Fig.  105,  and  is  collected  at  a  lower  level  in  the  boot  w,  Fig.  102, 
of  the  mercury  elevator,  shown  at  the  right.  The  elevator  discharges 
to  a  mercury  tank  s  commanding  the  amalgamating  pans,  to  which 
it  is  delivered  as  needed  through  pipe  shown  in  the  figure.  Over 
the  stamps  and  the  pans  are  seen  the  overhead  tracks  that  carry 


OF    THE    COMMON    METALS.  227 

crawls  by  which  the  heavy  parts  of  the  machines  are  lifted  or 
transferred.  This  facilitates  the  work  of  repairs  and  replacements. 

The  loss  of  mercury  is  commonly  1  to  1.5  Ib.  per  ton  of  ore 
treated.  A  part  is  lost  in  handling,  but  the  principal  cause  of 
loss  is  the  flouring  which  causes  the  mercury  to  escape  in  the 
tailings.  The  loss  is  greater  with  talcose  or  clayey  ores,  and  in 
those  carrying  cerussite,  chalcopyrite,  or  galena.  Loss  is  caused 
by  grease  coating  the  particles  of  mercury,  in  case  this  enters 
the  ore  from  the  machinery. 

Treatment  of  the  amalgam. — Since  the  weight  of  metal  recovered 
in  silver  milling  is  much  greater  than  in.  gold  milling,  the  retorting 


Fig.    105.     AMALGAM  SAFE. 

of  amalgam  must  be  performed  on  a  larger  scale.  Fig.  106  shows 
a  sectional  elevation  and  a  plan  of  a  combined  retorting  and  melting 
furnace  with  the  overhead  crawl  and  chain-blocks  by  which  the 
large  melting  crucibles  are  lifted  from  the  fire  in  the  melting 
furnace  and  transferred  for  pouring.  At  the  left  is  shown  in  the 
elevation  a  cross-section  of  the  cast-iron  cylindrical  retort  which  is 
10  in.  diam.  inside  by  28  in.  long,  resting  upon  arched  cast-iron 
supports.  There  is  a  horizontal  pipe,  and,  not  shown  in  the 
illustration,  a  vertical  water-cooled  pipe  in  which  the  mercury 
.condenses  and  from  which  it  falls  into  a  tub  of  water  below.  As 
seen  in  the  plan,  the  front  end  of  the  retort  is  provided  with  a 
cover  which  can  be  securely  clamped  in  position. 


228 


THE    METALLURGY 


The  charge  of  amalgam,  containing  20%  mercury,  should  weigh 
500  Ib.  and  only  half  fill  the  retort.  After  filling,  the  cover  is 
clamped  on,  first  luting  the  joint  with  flour  paste.  A  wood  fire 
is  started  on  the  grate  under  the  retort.  The  temperature  is  kept 


Fig.    106.      HORIZONTAL   RETORT   AND   MELTING   FURNACE   FOR 
SILVER  MILL. 

low  at  first,  increasing  to  a  red-heat  at  the  end,  %  to  %  cords 
of  wood  being  used.  The  operation  lasts  10  to  14  hours,  care  being 
taken  not  to  heat  the  retort  rapidly,  nor,  for  fear  of  blistering 
it,  to  raise  the  temperature  too  high.  The  fire  is  then  allowed 


OF   THE    COMMON    METALS.  229 

to  burn  down,  and  the  retort  to  cool.     The  lid  is  taken  off  and 
the  silver  residue  removed. 

The  melting  of  the  residue  is  performed  in  plumbago  crucibles, 
adding  soda  and  borax  as  fluxes.  The  crucible  is  set  in  the  wind- 
furnace  (shown  in  Fig.  106)  at  the  right  and  supported  upon  a 
brick  resting  on  the  grate-bars.  It  is  surrounded  with  coke  which 
burns  by  natural  draft,  the  smoke  escaping  by  the  flue  at  the 
back  to  the  stack.  When  the  charge  is  melted  the  crucible  and 
contents  are  removed  from  the  fire  with  basket-tongs,  which  fit  the 
crucible  and  clasp  it  firmly  so  that  it  can  be  lifted  by  the  chain- 
hoist,  transferred  by  the  crawl  to  the  ingot,  and  poured.  These 
molds,  11  in.  long  by  4%  in.  wide  and  deep,  hold  1000  oz.,  or  70  lb., 
silver. 

The  settler  tailing  contains  heavy  unaltered  ore  which  it  may 
pay  to  concentrate  on  a  Frue  vanner  of  some  similar  table.  At  some 
places,  sluices  are  used,  two  or  three  in  parallel.  These  may  be 
2  in.  high,  20  in.  broad,  and  1800  ft.  long.  The  bottoms  of  the 
sluices  are  covered  with  coarse  blanketing  which  can  be  easily 
removed  and  washed  to  recover  the  concentrate  or  particles  of 
amalgam  that  settle  into  the  meshes  of  the  blankets. 

Costs. — The  cost  of  pan-amalgamation  with  tank-settling 
( Washoe  process)  per  ton  of  ore  treated  is : 

Power     $0.087 

Labor    0.361 

Chemicals   (salt,   acid,  bluestone) 0.465 

Loss  of  mercury    0.750 

Wear  of  pans   0.200 

Wear  of  dies  and  shoes 0.400 

Oil,  interest,  and  superintendence 0.100 


Total    cost   per   ton $2.363 

One  notes  in  particular  the  larger  cost  of  supplies  (chemicals, 
mercury,  and  castings)  compared  with  like  items  in  gold  milling. 

55.     THE  BOSS  PROCESS  OF  SILVER  MILLING. 

This  system,  originated  by  M.  P.  Boss,  a  California  engineer, 
differs  from  the  Washoe  process  in  being  continuous  and  generally 
requiring  less  labor.  However,  the  Allis-Chalmers  Co.,  have 
designed — for  the  Washoe  process — a  wet-crushing  mill  in  which 
the  settling-boxes  have  sloping  bottoms,  so  arranged  that  the  content 
is  transferred  to  the  pans  with  but  little  labor.  This  takes  away 


230 


THE    METALLURGY 


the  advantage  urged  in  favor  of  the  Boss  system.  It  may  be  added 
that  the  settling  of  the  pulp  in  large  tanks,  combined  with  a 
mechanical  system  of  excavating  the  content  as  in  the  cyanide 
process,  ought  to  be  efficient  and  labor-saving. 

The  Boss  system  may  be  applied  to  free-milling  ores  and  to 
rebellious  ores  that  need  to  be  first  roasted. 

Fig.  107  is  a  sectional  elevation  and  Fig.  108  a  plan  of  a  30-stamp 
mill  arranged  according  to  the  Boss  system.  Fig.  108  shows  that  the 
ore,  coming  in  by  two  tram-tracks,  is  dumped  over  grizzlies,  the 
fine  falling  into  the  storage  bins,  while  the  coarse  ore  falls  upon  a 


Cross  Section 


Fig.   107      BOSS-PROCESS  SILVER  MILL    (ELEVATION). 

feed-floor  between  them.  Here  the  coarse  is  crushed  in  a  Blake 
crusher,  discharging  to  the  storage  bin  below.  From  the  bins 
the  ore  is  fed  automatically  to  the  stamps,  and  wet-crushed.  The 
screen-discharge  from  each  ten  stamps  flows  through  pipes  to  a 
pair  of  4-ft.  grinding-pans  in  series  or  six  pans  in  all  (shown  in 
the  plan,  Fig.  108)  where  the  ore  is  finely  ground.  It  overflows 
from  the  top  of  the  second  pan  into  a  common  launder  and  is 
carried  by  pipe  to  the  next  series  of  pans  at  a  lower  level,  flowing 
continuously  through  the  line  of  pans  and  settlers  there  situated. 
There  are  ten  amalgamating  pans  and  four  settlers.  They  are 
connected  one  to  another  near  the  top  by  4-in.  nipples,  which 
permit  the  flow  of  a  pulp  that  is  thinner  than  that  in  the  tank 
system. 


232  THE    METALLURGY 

The  amalgamating  pans  are  provided  with  double  bottoms  for 
the  admission  of  exhaust  steam  from  the  engine,  to  heat  the  pulp 
as  in  the  tank  system.  The  chemicals  are  fed  continuously  to 
the  first  two  pans  in  the  series  by  two  'chemical-feeders,'  and  the 
mercury  also  is  supplied  continuously  to  all  the  pans  by  pipes 
leading  from  the  mercury  distributing  tank.  Whatever  mercury 
settles  in  any  of  the  amalgamating  pans,  or  in  the  settlers,  is 
collected  in  the  mercury  well  of  each,  and  the  overflow  from  all 
the  wells  flows  through  suitably  arranged  pipes  to  the  strainer. 
The  strained  mercury  is  elevated  to  the  distributing  tank  to  be 
again  used.  The  continuous  flow  from  the  last  settler  goes  to  three 
settling  cones  in  series.  The  spigot-discharge  from  each  cone  goes 
to  a  separate  table,  and  the  overflow  to  a  fourth  one.  The  fifth 
table  is  used  for  re-treating  the  heads  of  the  first  four. 

Steam  siphons  are  provided  for  cleaning  out  the  pans,  or  for 
carrying  the  pulp  past  any  pan  when  necessary  to  cut  out  the  pan 
for  repairs.  The  main  line  shaft  runs  directly  under  the  pans 
and  s.ettlers,  each  of  which  is  driven  by  a  friction  clutch.  Thus 
any  machine  may  be  stopped  in  case  of  an  accident,  or  for  cleaning, 
without  having  to  stop  the  whole  line.  All  the  battery-flow, 
including  the  slime,  must  pass  through  the  system,  and  no  ore 
passes  untreated  as  in  the  tank  system.  The  loss  of  mercury  is 
small  and  it  is  possible  to  correctly  sample  the  tailing,  which 
are  advantages  over  the  older  system.  In  spite  of  the  thin  pulp 
used  it  is  claimed  that  an  ore  adopted  to  pan-treatment  can  be 
more  successfully  worked  by  the  continuous  system. 

56.     THE    COMBINATION    PROCESS    OF    SILVER    MILLING. 

This  process  is  used  on  ores  carrying  silver,  gold,  and  sulphides 
of  the  heavy  metals,  such  as  galena,  blende,  and  pyrite,  and 
sulphides  which  contain  silver  and  gold.  It  is  necessary  that  the 
silver,  not  in  the  sulphides,  be  amalgamable,  as  is  silver  chloride, 
argentite,  or  native  silver. 

The  process  consists  in  wet  stamping  the  ore,  running  the  pulp 
over  apron  plates  as  in  gold  milling,  concentrating  the  sulphides, 
and,  as  in  the  Washoe  process,  pan-amalgamating  the  tailing  and 
treating  the  amalgam  to  recover  the  silver  and  gold. 

Compared  with  either  wet  or  dry  silver-milling  the  process  has 
much  to  commend  it.  The  ore  being  refractory,  the  wet  process 
would  recover  little  value.  The  tonnage  stamped  by  the  dry-method 
with  roasting  would  be  low  compared  with  wet-stamping  which  is 
one-and-one-half  to  twice  as  rapid.  It  is  true  that  by  dry-stamping 


234  THE    METALLURGY 

and  roasting  we  are  able  to  extract  at  least  10%  more  metal  than 
can  be  obtained  by  raw  amalgamation,  but  this  is  offset  by  the 
cost  of  treatment  and  the  loss  of  precious  metal  in  roasting.  The 
combination  process  also  saves  lead  and  removes  galena,  sulph- 
arsenides,  and  sulph-antimonides,  all  of  which  tend  to  foul  and 
cause  the  loss  of  mercury.  Such  minerals  are  not  amenable  to 
amalgamation,  and  by  removing  them  for  separate  treatment  there 
results  a  cleaner  or  higher-grade  bullion.  Manganese  minerals  that 
consume  chemicals  in  the  pan  are  also  removed  by  concentration. 
Because  of  the  advantages,  it  is  probable  that  the  field  of  the 
combination  process  will  extend.  Certainly  it  will  be  used  in 
preference  to  roasting  and  amalgamation,  a  method  rapidly  losing 
ground ;  and  in  some  instances  it  may  take  the  place  of  lixiviation, 
so  far  as  respects  the  pan-amalgamation  feature  of  the  treatment. 

The  combination  mill. — Fig.  109  is  a  perspective  view  of  a 
10-stamp  combination  mill.  It  shows  at  the  left  a  tram-car  on 
the  upper  stage  which  is  about  to  be  emptied  over  the  grizzly. 
The  fine  falls  through  the  grizzly  while  the  lumps  fall  upon  the 
feed-floor  and  are  shoveled  into  a  Blake  crusher,  crushed  to  1-in. 
size,  and  united  with  the  fine  in  the  inclined  bottom  storage-bin 
below. 

Let  us  suppose  we  are  to  treat  an  ore,  in  part  oxidized,  but 
containing  the  heavy  minerals  of  lead  and  copper,  with  p3^rite, 
arsenides,  and  manganese  minerals.  The  ore  contains  the  precious 
metals,  a  gangue  of  quartz,  calcite,  and  a  little  clay,  and 
disseminated  through  it  gold  and  the  amalgamable  silver  minerals 
cerargyrite,  argentite,  and  native  silver.  The  purpose  is  to  save 
the  precious  metals  by  plate  and  pan-amalgamation,  and  the  heavy 
minerals  with  silver  and  gold  by  concentration.  Some  of  the  silver 
and  gold  escapes  recovery  and  is  lost  in  the  tailing.  Since  sulph- 
arsenides  and  manganese  minerals  are  mostly  removed,  they  do 
not  interfere  with  subsequent  pan-amalgamation  where  arsenic 
would  sicken  the  mercury  and  manganese  consume  chemicals. 

The  ore  and  water  are  fed  automatically  to  a  ten-stamp  battery, 
each  stamp  crushing  4  tons  per  24  hours  to  pass  a  30-mesh  screen. 
The  pulp  issuing  from  the  mortar  flows  over  two  apron-plates  (one 
for  each  5-stamp  mortar)  and  a  part  of  the  gold  and  silver  is 
recovered.  The  flow  is  distributed  evenly  to  four  Frue  vanners 
at  a  lower  level,  the  concentrate  (10%  of  the  whole)  being  separated 
to  ship  to  smelting  works,  while  the  tailing  is  carried  to  the  ten 
settling  boxes  in  a  double  row.  These  are  seen  at  the  left  of 
the  pans.  The  distribution  is  into  a  double  launder  between  the 


OF    THE    COMMON    METALS.  235 

two  rows.  By  drawing  the  plugs  in  the  bottom  of  the  launder,  the 
flow  can  be  directed  into  any  box  desired.  From  this  point  on,  the 
operation  is  conducted  as  described  for  the  Washoe  process.  There 
are  four  amalgamating  pans  and  two  settlers.  Bluestone  and  salt 
are  used  to  decompose  the  argentite.  Mercury  or  amalgam  escaping 
the  apron-plates  finds  its  way  into  the  settling  boxes  and  thence 
to  the  pans,  and  is  more  thoroughly  recovered  than  would  be 
possible  if  the  recovery  depended  upon  obtaining  it  in  the  vanner 
concentrate  as  in  gold  milling.  The  four  5-ft.  combination  grinding 
and  amalgamating  pans  each  treat  3000  Ib.  per  charge,  and  with 
a  4-hour  treatment,  this  equals  36  tons  daily,  which  with  the  4 
tons  of  concentrate  already  mentioned  is  a  40-ton  output  of  the 
mill.  Some  ores,  not  so  readily  treated,  take  6  to  8  hours,  and  lessen 
the  capacity  of  the  mill  accordingly. 

The  recovery  of  precious  metals  in  a  certain  ore,  containing 
0.40  oz.  Au  and  9  oz.  Ag  per  ton,  was  as  follows,  stated  in 
percentages : 

Gold        Silver 

Recovered  on  the  apron  plates 22  3 

Recovered  on  the  Frue  vanners 28  32 

Recovered  in  amalgamating  pans.  . .  ,32  35 

-Lost  in  the  tailing 18  30 


100  100 

The  total  recovery  of  the  gold  was  accordingly  82%,  and  of 
the  silver  70%.  Of  the  lead  and  copper  85%  was  saved  in  the 
concentrate.  The  tailing  contained  0.1  oz.  Au  and  3  oz.  Ag  per 
ton.  Concentrating  adds  but  little  to  the  cost  of  the  pan 
amalgamation,  and  $3  per  ton  may  be  taken  as  a  fair  estimate  of 
the  cost  of  combination  milling. 

57.     CHLORIDIZING  ROASTING  OF  SILVER  ORES. 

Silver  ore  containing  sulph-arsenides,  sulph-antimonides,  or 
tetrahedrite,  cannot  be  treated  directly  by  amalgamation  nor  by 
any  of  the  lixiviation  methods.  Such  ore  must  be  subjected  to  a 
roast  with  salt  to  convert  the  silver  into  a  chloride,  before  it  can 
be  successfully  treated  by  any  of  these  methods.  The  above  minerals 
are  often  accompanied  by  pyrite,  blende,  chalcopyrite,  and  galena. 

Preliminary  to  roasting,  such  ore  is  dry-crushed,  either  by 
rolls  or  by  stamps.  Ores  containing  galena  and  blende  are 
preferably  crushed  to  40-mesh  size,  those  having  pyrite  to  8  to  10 
mesh.  The  roasting  is  done  in  a  reverberatory  furnace,  and  requires 


236  THE    METALLURGY 

the  use  of  salt.  There  must  also  be  3  to  8%  pyrite  present  to 
furnish  sulphur  for  the  reactions,  and  if  the  ore  does  not  contain 
this,  it  must  be  added.  If  more  than  8%  sulphur  is  present,  the 
percentage  is  reduced  to  that  point  by  roasting  afterward,  before  the 
salt  is  added.  The  amount  of  salt  required  varies  according  to 
the  quantity  of  copper  and  iron  sulphides  present  which  consume 
the  evolving  chlorine. 

The  roasting  operation  is  at  first  an  oxidizing  one  conducted 
at  the  temperatures  specified  in  the  chapter  on  roasting.  The 
action  is  chiefly  upon  the  heavy  metals,  converting  them  into  either 
oxides  or  sulphates.  It  may  be  divided  into  three  stages  (1)  the 
kindling,  (2)  the  desulphurization,  and  (3)  the  chlorination  of 
the  ore. 

In  the  first  or  kindling  stage  we  find  the  loosely  held  sulphur 
being  driven  off,  and  the  ore  taking  fire,  producing  a  blue  flame. 

In  the  second  stage,  the  air  oxidizes  the  sulphides,  and 
particularly  the  newly  formed  iron  sulphide.  Reacting  upon  the 
sulph-antimonides  and  arsenides,  it  volatilizes  them  and  removes 
them  from  the  ore.  Copper  and  iron  sulphates  are  also  formed, 
the  later  according  to  the  following  reaction: 

3FeS  +  110  =  2SO2  +  Fe,03  +  FeS04 

In  the  third  stage,  at  590°  C.,  the  suiphate  formed,  reacts  upon 
the  salt,  thus : 

2FeS04  +  2NaCl  =  Na2S04  +  2FeO  +  S02  +  2C1 

The  odor  of  sulphur  dioxide  and  chlorine  is  pronounced.  The 
chlorine  thus  liberated  acts  at  once  upon  the  silver  compounds 
and  converts  them  into  chlorides. 

Zinc  blende  becomes  oxide  and  zinc  sulphate,  while  sulphur 
dioxide  escapes.  Galena  and  zinc  sulphate  remain  inactive  and 
fail  to  decompose  the  salt.  They  roast  slowly,  while  pyrite,  in 
presence  of  salt,  decomposes  quickly,  and  generates  chlorine  at  a 
period  in  the  roasting  when  neither  the  blende  nor  galena  are 
sufficiently  oxidized  to  expose  silver  to  the  action  of  the  chlorine. 
If  therefore  the  salt  is  mixed  with  the  ore  at  the  battery,  the 
chlorine  generated  by  the  reaction  of  the  ferrous  sulphate  and 
salt  is  lost,  an  imperfect  chlorination  results  no  matter  how  long 
roasting  is  continued,  nor  how  much  salt  is  added.  Hence  in 
roasting  an  ore  containing  blende  and  galena  it  is  of  the  greatest 
importance  to  add  the  salt  later  and  not  at  the  battery.  On  the 
other  hand  if  the  roasting  continues  until  the  sulphides  are  well 
oxidized,  the  iron  sulphate  decomposes  and  no  chlorine  is  generated 


OF    THE    COMMON    METALS.  237 

and  again  we  have  a  badly  chloridized  ore.  The  desirable  time 
to  add  the  salt  is  after  continued  roasting  at  a  low  heat  that  does 
not  break  up  the  iron  sulphate.  This  is  shown  when  the  black 
color  of  the  ore  changes  to  brown,  but  shows  still  the  presence  of 
black  particles.  A  distinct  odor  of  chlorine  is  then  to  be  noticed, 
due  to  the  decomposition  of  the  salt.  The  best  results  could  be 
obtained  by  adding  a  mixture  of  green  vitriol  (ferrous  sulphate) 
and  salt;  but  the  ore  would  hardly  justify  the  expense.  For 
moderate  quantities  of  ore,  the  chloridizing  roast  is  performed  in 
reverberatory  furnaces.  For  large  quantities,  roasting  is  cheaper 
when  conducted  in  the  mechanical  roasters. 

The  salt  is  added  to  the  dry  ore  at  the  time  of  charging,  if 
the  percentage  of  sulphur  is  suitable,  or  later  if  the  excess  of 
sulphur  must  be  first  removed  by  roasting.  The  temperature  is 
increased  only  gradually  to  kindle  or  start  the  ore  to  burning  and 
to  begin  oxidation.  As  the  temperature  rises  oxidation  and  the 
formation  of  sulphates  occur,  and  at  the  necessary  high  temperature 
these  act  upon  and  decompose  the  salt  and  chloridize  the  ore. 

It  is  not  considered  necessary  to  continue  the  roasting  to 
convert  all  possible  silver  into  chloride,  but  to  withdraw  the  charge 
while  hot  before  this  stage  is  reached.  During  the  gradual  cooling 
(12  to  30  hours)  further  chloridizing  proceeds,  due  to  the  action 
of  the  free  chlorine,  with  which  the  ore  is  saturated,  acting  on 
the  undecomposed  silver  sulphide.  This  may  increase  the  cloridiza- 
tion  10  to  40  per  cent. 

Upon  completion  of  the  operation  of  'heap  chlorination',  as  it  is 
called,  and  with  ores  containing  copper  chloride,  a  wetting  down  or 
sprinkling  causes  an  additional  chlorination  of  3  to  6%.  Thus  at 
the  Lexington  mill,  Butte,  Montana,  the  ore,  after  roasting  in  a 
Stetefeldt  furnace  was  chloridized  to  65%  after  two  hours  in  the 
heap  to  75  or  80%,  and  at  the  end  of  36  hours  to  92%  of  the  silver 
content. 

The  loss  in  silver  by  volatilization,  when  the  ore  has  been 
properly  and  carefully  roasted,  should  not  exceed  8%  except  in 
presence  of  volatile  elements  like  arsenic,  antimony,  selenium,  or 
tellurium.  If,  however,  the  roasting  is  completed  at  a  high  temper- 
ature the  loss  may  rise  to  18  per  cent. 

The  most  difficult,  and  at  the  same  time  the  most  important 
process  for  the  treatment  of  silver  ores  by  wet-methods,  is 
undoubtedly  chloridizing  roasting.  It  is  always  the  safest  plan  for 
the  operator  to  roast  as  thoroughly  as  possible.  If  the  ore  is  well 
chloridized  sodium  hyposulphite  extracts  all  the  silver  chloride. 


238  THE    METALLURGY 

A  high  chloridization  does  not  necessarily  involve  a  high  loss  by 
volatilization. 

58.     DRY  SILVER-MILLING  (REESE  RIVER  PROCESS). 

This  process  for  the  treatment  of  rebellious  silver  ores,  in  which 
the  metal  is  so  locked  up  as  to  require  roasting  before  it  can  be 
amalgamated,  was  developed  at  Reese  river,  near  the  Comstock 
Lode  at  Virginia  City,  Nevada.  The  ore  contains  silver  sulphide, 
particularly  the  antimonial  sulphides,  and  the  sulphide  of  the  base 
metals  such  as  copper,  iron,  zinc,  and  lead.  Galena,  however,  if 
present  exceeding  5  to  10%,  renders  the  ore  unsuitable  for 
chloridization. 

The  treatment  in  brief  consists  in  dry-crushing  and  roasting 
the  ore,  then  amalgamating  in  pans  to  recover  the  silver  and  gold. 
The  dry-crushing  is  done  either  with  rolls  or  stamps.  Crushing  with 
rolls  is  described  in  the  chapter  on  'Crushing.'  If  dry  stamping 
is  employed  the  work  is  done  in  the  dry-crushing  silver  mill,  Fig. 
110  and  111,  here  described. 

Fig.  110  is  a  sectional  elevation,  and  Fig.  Ill  a  plan  of  a  20-stamp 
mill.  The  ore  is  dumped  from  the  tram-car  over  the  grizzly  6,  which 
separates  the  fine  from  the  lumps  which  roll  to  the  mouth  of  the 
crusher  c  and  are  crushed  to  l^-in.  size.  The  crushed  product 
falls  with  the  fine  into  the  storage  bin  d.  From  this  the  ore  is 
fed  continuously  throughout  the  24  hours  by  means  of  an  automatic 
feeder  e,  which  supplies  a  revolving  drier  /  that  is  18  ft.  long.  The 
drier  is  provided  with  a  fire-box  g,  the  smoke  passing  through  a 
flue  to  a  separate  stack.  Flue-dust  is  retained  in  the  flue. 

The  ore  from  the  drier  containing  less  than  1%  moisture,  slides 
by  an  inclined  launder  or  feed-chute  to  the  automatic  feeders  i  of 
the  stamps.  The  fine  particles  drive  through  the  screens,  while 
the  coarse  part  drops  back  upon  the  die  to  be  further  crushed. 
Naturally  the  process  produces  dust,  and  to  prevent  escape  into 
the  mill,  the  mortar  is  housed  or  boxed.  There  is  also  an  exhaust-fan 
connected  with  the  housing  by  which  the  dust  is  drawn  to  a  dust- 
chamber  and  settled.  The  finely  ground  ore  passing  through  the 
screens,  drops  to  the  screw-conveyors,  one  at  the  front  of  the  other 
at  the  back  of  the  mortars,  and  is  conveyed  by  them  to  the  hopper 
or  boot  of  the  belt-elevator  fc. 

It  is  customary  to  feed  into  the  last  battery  the  rock  salt 
required  by  the  ore.  Being  finely  crushed  it  can  be  intimately 
mixed  with  the  ore.  The  elevator  raises  the  mixed  ore  and  salt  to 
the  conveyor  m.  Thence  the  screw  conveyor  n  delivers  it  into  the 


240  THE    METALLURGY 

hopper  of  the  roaster.  (In  Fig.  194  is  shown  the  manner  in  which 
the  transfer  from  ra  to  n  is  effected,  the  delivering  conveyor  being 
shown  at  the  right,  and  the  receiving  one  at  the  left.)  From  the 
feed-hopper  the  ore  slides  into  the  White-Howell  roasting  furnace 
o  (See  also  Fig.  34),  where  it  receives  the  chloridizing  roast. 

In  the  plan  of  the  mill,  Fig.  Ill,  is  to  be  seen  the  flue-chamber, 
and  a  stack  like  that  for  the  ore-dryer.  In  the  chamber  the  dust 
produced  in  roasting  settles  and  is  recovered.  The  roasted  ore, 
discharged  from  the  roaster,  is  received  into  one  of  the  hoppers  r. 
When  either  hopper  is  filled,  the  content  is  drawn  off,  moistened 
with  a  little  water  to  prevent  dusting,  and  stored  on  the  cooling 
floor  s.  The  level  of  the  floor  s  is  slightly  above  that  of  the  pan- 
floor  where  the  man  in  Fig.  110  is  seen  standing.  The  tracks  on  the 
pan-floor-level  are  arranged  so  that  a  tram-car  can  take  the  ore 
from  the  cooling  floor  to  any  of  the  ten  amalgamating  pans  t. 

The  amalgamation  of  roasted  ore  is  performed  in  pans  having 
wooden  sides  (See  Fig.  103.)  as  in  the  Washoe  process.  Water  is 
run  into  the  pan  first,  and  while  in  motion,  the  3000-lb.  charge  of 
ore  is  shoveled  in  until  the  pulp  is  of  the  consistence  of  honey.  If 
the  ore  is  imperfectly  roasted,  salt  and  bluestone  are  added  to 
decompose  the  silver  sulphide.  The  free  chlorine  in  the  ore  would 
be  consumed  by  contact  with  the  iron  surfaces  of  the  pan,  and  hence 
the  wooden  sides  are  provided.  After  a  time,  the  iron  particles  that 
have  come  from  the  stamps  are  thus  removed  and  prevented  from 
acting  upon  the  mercury  and  converting  it  to  mercury  chloride  in 
a  way  that  would  be  detrimental  to  the  process.  When  the  ore 
has  been  ground  two  hours  the  shoes  are  raised  about  half  an  inch 
from  the  dies,  and  300  Ib.  mercury  is  added.  If  mercury  were  added 
during  the  grinding  it  would  flour  seriously.  The  shoes  and  muller 
intimately  mix  the  pulp  and  mercury.  Amalgamation  proceeds 
rapidly  at  first,  but  more  slowly  toward  the  last,  the  silver  chloride 
being  reduced  to  metal  as  in  the  Washoe  process.  The  iron  present 
reduces  the  higher  chlorides  of  copper  and  iron,  and  precipitates 
mercury  from  the  mercurous  chloride.  The  mixing  is  of  six  hours 
duration.  Water  is  added  and  the  entire  content  of  the  pan  is 
run  into  the  settlers  u  Fig.  111.  Here  it  is  diluted  with  water  and 
the  mercury  and  amalgam  settled.  The  tailing  is  decanted,  and 
the  amalgam  is  recovered  and  treated  as  described  under  the  Washoe 
process. 

The  recovery  of  silver  is  thorough,  being  in  some  instances  97%. 
The  loss  of  quicksilver  is  slight,  being  less  than  i/±  Ib.  per  ton 
of  ore.  At  the  Lexington  mine,  ore  containing  28.5  oz.  Ag  and 


Fig.    111.      DRY-CRUSHING  SILVER  MILL    (PLAN). 


242  THE    METALLURGY 

0.58  oz.  Au  per  ton  yielded  93.3%  of  the  silver  and  60%  of  the 
gold  after  roasting.  The  loss  of  silver  in  roasting,  however,  was 
1%,  and  that  of  gold  20%.  This  gives  in  actual  recovery,  86.8% 
of  the  silver,  but  only  48%  of  the  gold. 

The  cost  of  dry  stamping  followed  by  chloridizing  roasting  may 
be  considered  $6.48  per  ton  of  ore  treated,  this  being  an  average 
of  results  in  three  mills  that  vary  but  little  from  one  another, 
situated  in  different  parts  of  the  Rocky  Mountain  country. 

Upon  comparing  the  recovery  at  the  Lexington  mine  with  one  of 
60%  of  the  silver  and  gold  as  obtains  by  direct  amalgamation  of 
the  raw  ore,  we  have,  with  the  price  of  silver  60c.  and  of  gold 
$20.65  per  ounce,  the  following  result: 

Raw        Roasted 

Treatment     $2.46  $6.48 

Loss  of  silver 6.84  2.14 

Loss  of  gold 4.79  4.79 


$14.09  $13.41 

It  is  seen  in  the  extraction  represented,  that  there  is  but  little 
advantage  in  roasting  over  the  raw  treatment  of  ore  of  this  grade, 
to  say  nothing  of  the  extra  cost  of  installing  a  roasting  plant,  and 
the  large  mill  needed  for  the  tonnage.  Of  course,  in  case  ore  is  so 
refractory  that  only  a  low  extraction  of  metal  is  possible,  a  chloridiz- 
ing roast  may  be  the  alternative. 

While  rebellious  ores  have  been  worked  by  dry  silver-milling, 
it  pays  in  many  cases,  where  accessible  to  smelting-works,  to  ship 
ore  rather  than  to  mill  it.  If  the  heavy  part  of  the  ore  carries 
most  of  the  value,  it  is  better  to  concentrate  it  and  submit  to  the 
loss  in  the  tailing  to  gain  the  advantages  of  treating  a  concentrated 
ore.  Such  problems  are  solved  on  a  trial  scale  before  undertaking 
the  erection  of  a  plant. 

Comparing  dry  silver-milling  with  smelting,  where  the  smelter 
pays  for  95%  of  silver  and  for  the  gold  at  $19.50  per  ounce,  and 
makes  a  freight  rate  of  $5  and  a  treatment  rate  of  $8  per  ton  we 

have  the  following  condition : 

Smelting      Dry  Silver-Milling 

Treatment $8.00  $6.48 

Freight    5.00 

Loss  in  silver 0.86  2.14 

Loss  in  gold 0.67  4.79 


$14.53  $13.41 


OF    THE    COMMON    METALS.  243 

This  indicates  a  gain  of  $1.12  in  favor  of  roasting-amalgamation 
over  smelting. 

On  the  other  hand  the  cost  of  erecting  an  expensive  plant  is 
saved  by  shipping  the  ore  to  the  smelting-works  as  is  today 
generally  done. 

59.     SILVER  MILLING  AT  BLACK  PINE,  NEVADA. 

The  ore  of  the  Combination  Mining  &  Milling  Co.,  at  Black 
Pine,  Nevada,  contains  silver  in  malachite  and  tetrahedrite,  with 
a  quartz  gangue,  and  no  lead  or  zinc  minerals.  The  process 
consists  in  milling  and  concentrating  the  ore  to  separate  the 
refractory  tetrahedrite  from  the  free-milling  malachite  and  roasting 
and  pan-amalgamating  the  concentrate. 

The  ore  is  partly  sorted  underground,  and  upon  hoisting,  dumped 
over  grizzlies  in  a  rock-house,  the  fine  falling  into  a  hopper-bottom 
bin,  to  be  thence  hauled  by  team  to  a  sloping  bottom  storage-bin 
near  the  mill  in  which  a  good  supply  can  be  carried.  From  this 
bin  the  ore  is  trammed  to  the  mill,  and  dumped  over  a  grizzly  with 
one-inch  openings  to  separate  the  fine.  The  coarse  ore  falls  to  the 
mouth  of  a  10  by  20-in.  Blake  crusher  running  250  rev.  per  minute. 
The  crushed  product  joins  the  fine  in  the  mill  storage-bin  as 
described  in  connection  with  Fig.  102.  All  the  foregoing  operations 
occur  in  the  day  shift. 

From  the  mill  storage-bin  the  ore  is  fed  by  automatic  feeder 
to  a  20-stamp  battery,  having  1100-lb.  stamps  dropping  6  to  9  in. 
100  times  per  minute.  It  is  crushed  through  30-mesh  brass-wire 
screens  at  the  rate  of  3.75  tons  per  stamp  in  24  hours.  The  battery- 
pulp  is  carried  by  launder  to  a  belt  elevator  which  raises  it  above 
the  level  of  the  Frue  vanners.  The  pulp  is  discharged  into  a 
distributing  box  and  thence  to  the  twelve  vanners  of  6*4  tons 
daily  capacity  each  of  75  tons  total  capacity.  With  a  concentration 
of  four  into  one,  there  is  obtained  nearly  20  tons  of  concentrate 
containing  the  refractory  tetrahedrite  needing  chloridizing  roasting, 
and  55  tons  of  free-milling  tailing  containing  malachite.  The 
tailing  is  settled  as  in  the  Washoe  process  (See  Fig.  102  and  109) 
in  a  series  of  18  tanks,  each  9.25  ft.  long,  6.5  ft.  wide,  and  2.75 
ft.  deep. 

The  settled  pulp  is  shoveled  from  the  tanks  to  the  floor 
adjoining.  It  is  fed  as  needed  into  14  wooden-sided,  6-ft. 
amalgamating  pans,  the  charge  consisting  of  3000  Ib.  tailing  and 
1000  Ib.  concentrate  that  has  received  a  chloridizing  roast  to  be 
described.  To  the  charge  is  added  300  Ib.  mercury  to  amalgamate 


244  THE    METALLURGY 

the  silver,  17  Ib.  salt  to  react  with  the  copper  salts  present  and 
decompose  silver  sulphide,  8  to  16  Ib.  iron  filings  to  decompose 
chlorides,  and  1  Ib.  lye  to  correct  any  acidity  of  the  ore.  The. 
iron  filings  react  as  follows: 

GAgCl  +  2Fe  +  3Hg  =  3Ag2Hg  +  Fe2Cl6 

3CuCl2  +  2Fe  +  3Hg  =  3Hg  +  Fe2Cln 

A 
Without  the  iron,  mercury  chloride   (HgCL)   would  be  formed  and 

lost  in  the  pan-tailing.  Silver  and  copper  chloride  are  accordingly 
decomposed  and  the  metals  amalgamated. 

The  amalgamation  is  completed  in  eight  hours,  then  the  content 
of  a  pan  is  run  into  an  8-ft.  settler  39  in.  deep,  of  which  there 
are  seven.  The  settler  is  run  at  20  rev.  per  min.  and  the  charge, 
now  diluted  with  water,  is  worked  off  in  four  hours,  one  settler 
treating  the  output  of  two  pans.  At  the  side  of  the  settler  near 
the  top  is  a  rectangular  overflow  notch  3  in.  deep,  and  at  distances 
of  8,  14,  21,  and  25  in.  down  are  3-in.  plugs.  The  pan,  having 
been  for  an  hour  in  motion,  decanting  separated  pulp  and  water, 
the  first  plug-hole  from  the  top  is  opened.  After  another  hour 
the  second  plug  is  withdrawn  and  decantation  continued  two  hours 
longer,  the  sand  in  the  settler  below  the  second  plug  remaining 
until  the  periodical  cleaning  out.  The  plug-holes  are  now  closed, 
and  the  pan  is  ready  for  another  charge.  The  exhausted  tailing 
is  conveyed  to  a  settling  pond  and  impounded  behind  a  crib-dam 
thrown  across  the  canyon  below  the  mill. 

The  mercury. — From  the  mercury-well  at  the  side  of  a  settler 
(See  Fig.  104),  the  mercury  overflows  to  an  amalgam  safe  (See 
Fig.  105).  After  straining  through  the  canvas  filter-sack  inside 
the  safe,  it  drains  to  the  foot  of  a  4-in.  belt  elevator  provided  with 
cast-iron  cup-buckets.  This  raises  and  delivers  it  to  a  mercury 
storage-tank  where  it  is  reserved  until  again  used.  This  arrange- 
ment of  settler,  safe,  elevator,  and  tank,  is  well  shown  in  Fig.  102. 

To  remove  mercury  chloride  which  causes  mercury  to  flour  and 
fail  to  coalesce,  one  pound  of  sodium  is  stirred  into  16  Ib.  mercury 
and  added  to  the  mercury  supply  every  two  days.  The  sodium 
reacts  as  follows: 

2Na  +  HgCL  =  2NaCl  +  Hg. 

The  NaCl  then  dissolves  in  the  water.  Eight  pounds  of  potassium 
cyanide  also  is  added  weekly,  the  cyanide  reacting  upon  the  mercury 
chloride  thus : 

HgCL,  +  2KCN  =  Hg(CN)2  -f  2KC1. 


OF    THE    COMMON    METALS.  245 

The  KC1  passes  into  solution.     The  mercury  cyanide  is  soluble  in 
the  mercury  and  adds  to  its  activity. 

Treatment  of  amalgam. — The  amalgam  removed  from  the  canvas 
strainers  is  often  dirty  and  must  be  sent  to  a  clean-up  pan  to 
remove  the  waste-matter  before  retorting.  It  is  placed  in  the 
pan  for  cleaning,  ground,  settled,  and  the  sand  and  dirty  water 
decanted  as  from  a  settler.  The  amalgam  thus  formed  is  again 
put  through  a  strainer,  and  the  cleaned  amalgam  sent  to  the 
retorts. 

Retorting. — From  75  tons  of  25-oz.  ore  there  comes  1600  oz.  silver 
or  12,800  oz.  coppery  bullion  125  fine.  This  corresponds  to  64,000 
oz.  or  4400  Ib.  amalgam  of  20%  Ag.  There  are  four  retorts,  each 
capable  of  holding  1600  oz.  amalgam.  The  amalgam  is  charged 
to  the  retorts,  Fig.  106,  the  covers  luted  on,  and  the  retorting 
conducted  in  8  to  10  hours.  A  jacket  (See  Fig.  54)  that  surrounds 
the  condensing  tube,  is  supplied  with  cold  water  under  pressure. 
The  mercury-fumes  thus  condensed  are  caught  in  a  tub  containing 
water. 

Melting  the  residue. — This  is  done  in  an  efficient  way  in  a  small 
reverberatory  furnace,  which  is  first  heated  to  a  bright-red  heat. 
The  furnace  is  charged  with  4400  Ib.  retort  residue,  upon  which 
when  melted  is  thrown  in  20  Ib.  each  of  borax  and  nitre  to  act 
as  a  flux.  The  bath  is  then  thoroughly  stirred  and  two  samples 
for  assay  taken  with  a  long-handled  spoon,  and  granulated  by 
pouring  into  water.  Brass  molds,  holding  100  Ib.  each,  placed 
close  together  on  a  carriage,  and  painted  with  a  mixture  of  lamp- 
black and  benzine,  are  brought  beneath  the  furnace  spout  and 
the  bullion  tapped  into  them,  the  carriage  being  moved  along  as 
the  molds  are  successively  filled. 

Roasting  the  concentrate. — The  concentrate  amounts  to  20  tons 
daily.  It  is  trammed  to  the  roaster-room  and  emptied  on  the  floor 
above  the  Bruckner  roaster.  The  charge  is  shoveled  into  the 
cylindrical  roaster,  which  is  revolved  15  minutes  for  the  ore  to 
dry.  Then  10%  salt  is  added,  and  if  there  is  not  sufficient  sulphide 
in  the  ore  sulphur  is  added  to  increase  this  to  5%.  The  charge  is 
now  fired  and  the  temperature  gradually  raised,  dense  grayish- 
white  arsenical  and  antimonial  fumes  being  evolved.  Next,  with 
increasing  heat,  S03  escapes,  while  the  presence  of  CuSO4  is 
indicated  by  the  blue  copper  flame.  When  this  grows  faint  the 
chloridizing  reactions  begin,  and  a  bright-red  heat  is  maintained  to 
the  end.  Both  roasting  and  amalgamation  are  here  simplified  by 


248  THE    METALLURGY 

(2)     Ag2S  +  CuCl2  =  2AgCl  +  CuS 

The  above  reaction  is  slow,  and  it  is  probable  that  the  cupric 
chloride  acts  directly  on  the  mercury  as  follows : 

(3)     2CuCl2  +  2Hg  =  Cu2Cl2  +  Hg2Cl2 
The  mercurous  chloride  is  lost. 

While  Cu2Cl2  is  insoluble  in  water  it  is  soluble  in  salt  solution, 
and  according  to  Laur  can  act  directly  on  silver  sulphide  with  the 
production  of  metallic  silver : 

(4)     Ag2S  +  Cu2Cl2  ==  CuS  +  CuCl2  +  2Ag 
61.     THE  HYDRO-METALLURGY  OF  SILVER. 

A  wet-process  for  the  recovery  from  the  ore  consists  in 
dissolving  the  metal  by  means  of  a  solvent  and  precipitating  from 
the  solution  in  a  convenient  form.  The  silver  compounds  which 
can  be  obtained  readily  in  solution  are  the  sulphate  and  the  chloride. 
In  cyanide  solution  argentite  is  readily  soluble,  while  ruby  silver, 
freislebenite,  and  stephanite,  are  sparingly  so,  though  readily  soluble 
in  mercurous  potassic  cyanide.  Silver  sulphate  is  soluble  in  hot 
water,  while  silver  potassic  chloride  is  dissolved  by  brine  solution 
or  by  sodium  hyposulphite  (thiosulphate).  From  the  aqueous 
solution  of  the  sulphate  silver  is  precipitated  by  metallic  copper; 
from  the  brine  solution  of  its  chloride  by  copper,  or  when  in  dilute 
solution,  by  zinc  iodide ;  from  the  hyposulphite  solution  by  sodium 
sulphide ;  and  from  the  cyanide  solution  by  metallic  zinc. 

There  are  four  well  known  wet-processes  for  the  extraction  of 
silver : 

(1)  The  Augustin  process,  based  upon  the  solubility  of  silver 
chloride  in  brine. 

(2)  The  Ziervogel  process  dependent  on  the  solubility  of  silver 
sulphate  in  hot  water. 

(3)  The  Patera  process  in  which   silver  chloride   dissolves  in 
a  solution  of  sodium  hyposulphite. 

(4)  The   cyanide  process  in   which   the   silver  minerals   above 
enumerated  dissolve  in  dilute  potassium  cyanide  solution. 

62.     THE  AUGUSTIN  PROCESS. 

This  has  been  used  for  the  extraction  of  silver  from  ore  and 
from  copper-bearing  matte,  obtained  as  a  product  of  smelting. 
At  Kosaka,  Japan,  ore  consisting  of  one-half  heavy  spar  and 
containing  10.5  oz.  silver  per  ton  is  thus  treated.  The  ore  is  crushed 
and  roasted  with  salt  in  a  furnace  6,  Fig.  112,  and  after  drawing 


OF    THE    COMMON    METALS. 


249 


from  the  furnace  and  moistening  on  the  cooling  floor,  contains 
80%  of  silver  in  the  form  of  chloride.  It  is  leached  with  a  hot 
18%  salt  solution  in  regular  leaching  vats  r.  The  leaching  is 
continued  until  a  polished  plate  of  copper  shows  no  precipitate  of 
silver  when  held  in  the  flowing  filtrate.  It  requires  0.66  tons  of 
brine  to  leach  a  ton  of  the  ore.  The  sand  is  washed  with  hot 
water,  and  the  tailing  rejected. 

The  brine  solution  from  the  vat  c  is  received  in  a  precipitating 
tank  d.  The  tank,  like  a  leaching-vat,  has  a  false  bottom.  Upon 
the  false  bottom  is  spread  a  2-in.  bed  of  bean-copper,  and  on  this 
rest  plates  of  copper  6  by  8  in.  by  1  in.  thick.  The  silver, 


Fig.    112.      FLOW-SHEET   OF  AUGUSTIN  PROCESS. 

precipitated  in  crystalline  form  upon  the  copper,  is  called  'cement- 
silver.'  At  a  lower  level  is  a  tank  e  containing  scrap-iron  where 
the  copper  in  solution  is  precipitated,  while  the  barren  brine  goes 
to  the  brine-sump.  It  is  there  brought  up  to  the  full  strength  and 
pumped  back  to  be  again  used.  The  cement-silver,  150  to  750  fine, 
is  removed  from  the  plates,  squeezed  in  a  screw-press  into  'cheeses' 
12  in.  diam.  by  3%  in.  thick,  dried,  and  refined  in  an  English 
cupelling-furnace  in  charges  of  150  Ib.  with  300  Ib.  of  lead  added 
to  each  charge.  The  refined  silver,  the  result  of  the  operation,  is 
melted  in  crucibles  and  cast  in  bars  of  1000  oz.  each,  985  fine. 

Treatment  of  matte  by  the  Augustin  process, — As  indicated  in 
the  diagram  (See  Fig.  112)  the  matte  after  a  preliminary  roasting 
is  smelted  to  a  higher-grade  matte  in  furnace  a.  It  is  then  crushed, 


248  THE    METALLURGY 

(2)     Ag2S  +  CuCl2  =  2AgCl  +  CuS 

The  above  reaction  is  slow,  and  it  is  probable  that  the  cnpric 
chloride  acts  directly  on  the  mercury  as  follows : 

(3)     2CuCl2  +  2Hg  =  Cu2Cl2  +  Hg2Cl2 
The  mercurous  chloride  is  lost. 

While  Cu2Cl2  is  insoluble  in  water  it  is  soluble  in  salt  solution, 
and  according  to  Laur  can  act  directly  on  silver  sulphide  with  the 
production  of  metallic  silver  : 

(4)     Ag2S  +  Cu2Cl2  =  CuS  +  CuCl2  +  2Ag 
61.     THE  HYDRO-METALLURGY  OF  SILVER. 

A  wet-process  for  the  recovery  from  the  ore  consists  in 
dissolving  the  metal  by  means  of  a  solvent  and  precipitating  from 
the  solution  in  a  convenient  form.  The  silver  compounds  which 
can  be  obtained  readily  in  solution  are  the  sulphate  and  the  chloride. 
In  cyanide  solution  argentite  is  readily  soluble,  while  ruby  silver, 
freislebenite,  and  stephanite,  are  sparingly  so,  though  readily  soluble 
in  mercurous  potassic  cyanide.  Silver  sulphate  is  soluble  in  hot 
water,  while  silver  potassic  chloride  is  dissolved  by  brine  solution 
or  by  sodium  hyposulphite  (thiosulphate).  From  the  aqueous 
solution  of  the  sulphate  silver  is  precipitated  by  metallic  copper; 
from  the  brine  solution  of  its  chloride  by  copper,  or  when  in  dilute 
solution,  by  zinc  iodide ;  from  the  hyposulphite  solution  by  sodium 
sulphide ;  and  from  the  cyanide  solution  by  metallic  zinc. 

There  are  four  well  known  wet-processes  for  the  extraction  of 
silver  : 

(1)  The  Augustin  process,  based  upon  the  solubility  of  silver 
chloride  in  brine. 

(2)  The  Ziervogel  process  dependent  on  the  solubility  of  silver 
sulphate  in  hot  water. 

(3)  The  Patera  process  in  which  silver  chloride  dissolves  in 
a  solution  of  sodium  hyposulphite. 

(4)  The   cyanide  process  in   which   the   silver  minerals   above 
enumerated  dissolve  in  dilute  potassium  cyanide  solution. 

62.     THE  AUGUSTIN  PROCESS. 

This  has  been  used  for  the  extraction  of  silver  from  ore  and 
from  copper-bearing  matte,  obtained  as  a  product  of  smelting. 
At  Kosaka,  Japan,  ore  consisting  of  one-half  heavy  spar  and 
containing  10.5  oz.  silver  per  ton  is  thus  treated.  The  ore  is  crushed 
and  roasted  with  salt  in  a  furnace  &,  Fig.  112,  and  after  drawing 


OF    THE    COMMON    METALS. 


249 


from  the  furnace  and  moistening  on  the  cooling  floor,  contains 
80%  of  silver  in  the  form  of  chloride.  It  is  leached  with  a  hot 
18%  salt  solution  in  regular  leaching  vats  c.  The  leaching  is 
continued  until  a  polished  plate  of  copper  shows  no  precipitate  of 
silver  when  held  in  the  flowing  filtrate.  It  requires  0.66  tons  of 
brine  to  leach  a  ton  of  the  ore.  The  sand  is  washed  with  hot 
water,  and  the  tailing  rejected. 

The  brine  solution  from  the  vat  c  is  received  in  a  precipitating 
tank  d.  The  tank,  like  a  leaching-vat,  has  a  false  bottom.  Upon 
the  false  bottom  is  spread  a  2-in.  bed  of  bean-copper,  and  on  this 
rest  plates  of  copper  6  by  8  in.  by  1  in.  thick.  The  silver, 


\ 


Fig.    112.      FLOW-SHEET   OF  AUGUSTIN   PROCESS. 

precipitated  in  crystalline  form  upon  the  copper,  is  called  'cement- 
silver.'  At  a  lower  level  is  a  tank  e  containing  scrap-iron  where 
the  copper  in  solution  is  precipitated,  while  the  barren  brine  goes 
to  the  brine-sump.  It  is  there  brought  up  to  the  full  strength  and 
pumped  back  to  be  again  used.  The  cement-silver,  150  to  750  fine, 
is  removed  from  the  plates,  squeezed  in  a  screw-press  into  'cheeses' 
12  in.  diam.  by  3l/2  in.  thick,  dried,  and  refined  in  an  English 
cupelling-furnace  in  charges  of  150  Ib.  with  300  Ib.  of  lead  added 
to  each  charge.  The  refined  silver,  the  result  of  the  operation,  is 
melted  in  crucibles  and  cast  in  bars  of  1000  oz.  each,  985  fine. 

Treatment  of  matte  by  the  Augnstin  process, — As  indicated  in 
the  diagram  (See  Fig.  112)  the  matte  after  a  preliminary  roasting 
is  smelted  to  a  higher-grade  matte  in  furnace  a.  It  is  then  crushed, 


250 


THE    METALLURGY 


and  roasted  with  salt  in  the  furnace  6,  and  the  roasted  product 
leached  in  the  tank  c,  as  already  described  for  ore.  The  residue 
from  c  cannot  be  rejected,  as  was  the  case  with  the  ore,  for  it 
contains  copper  and  iron  oxides,  and  must  be  further  treated  b,y 
the  Welsh  method  as  described  in  the  chapter  on  the  metallurgy 
of  copper.  It  is  smelted  in  furnaces  /  with  silicious  ore  and  copper 
sulphates,  producing  a  copper  matte.  The  matte,  in  coarse  pieces, 
is  charged  into  another  furnace  from  the  vat  e,  and  slowly  melted 
by  an  oxidizing  flame.  The  smelting  is  completed  at  a  high 
temperature  producing  blister  copper.  This  is  cast  in  ingots  or 
pigs,  and  refined  in  a  furnace  to  produce  market  copper. 

63.     THE  ZIERVOGEL  PROCESS. 

This  process,  practised  at  Mansfeldt,  Germany,  and  at  the  Boston 
and  Colorado  smelting  works  at  Argo,  Colorado,  is  adopted  to 
the  treatment  of  rich  copper  matte  containing  little  or  no  arsenic, 


COPPER  ORE-S 

&  GOLD  ORES 


ORES 


FINAL  SOLUTION 
MARKET  COPPER  To  WASTE 

Fig.    113.      FLOW-SHEET   OF  ZIERVOGEL  PROCESS. 

antimony,  or  bismuth,  any  of  which  would  form  insoluble  compounds 
with  silver.  The  method  may  be  divided  into  three  parts :  the 
roasting  for  silver  sulphate,  the  leaching,  and  the  precipitation  of 
the  silver. 


OF    THE    COMMON    METALS.  251 

The  process. — Referring  to  the  flow-sheet  of  the  process  (See 
Fig.  113)  we  have  in  furnaces  a,  the  operation  of  producing  the 
matte  or  regulus  from  gold  and  silver-bearing  copper  ores.  The 
details  of  the  process  are  described  in  the  chapter  on  the  metallurgy 
of  copper,  under  the  head  of  ' Reverberatory  Matte  Smelting.'  The 
composition  of  the  matte  is  Cu,  47.3;  Pb,  8.1;  Zn,  2.7;  Fe,  17.7; 
S,  21.6%,  with  400  oz.  silver  and  15  oz.  gold  per  ton. 

Preparation  of  the  matte. — The  matte  is  crushed  and  passed 
through  rolls  at  1)  to  reduce  it  to  6-mesh  size,  and  sent  to  a  Pearce- 
turret  furnace  c  (See  also  Fig.  40  and  41),  where  it  receives 
preliminary  roasting.  The  roasting  reduces  the  sulphur  to  6.3%, 
and  converts  the  iron  and  copper  sulphides  to  the  corresponding 
oxides  and  sulphates,  as  described  in  the  chapter  on  the  chemistry 
of  oxidizing  roasting.  This  partly  roasted  product  then  goes  to 
a  Chilian  mill  d  (See  also  Fig.  100),  where  it  is  finely  ground  to 
60-mesh. 

Sulphatizing  roasting. — The  partly  roasted  matte  is  next  treated 
in  charges  of  1600  Ib.  by  a  sulphatizing  roast  in  small  single-hearth 
reverberatory  roasters  at  e.  In  the  process  the  iron  and  copper 
remaining  in  the  form  of  sulphides  are  converted  into  sulphates 
which  react  on  the  silver  sulphide  at  a  slightly  higher  temperature 
as  follows : 

Ag2S  +  30  +  CuS04  =  Ag2S04  +  CuO  +  S02 

It  has  been  found  that  the  addition  of  2%  sodium  sulphate  (salt 
cake)  facilitates  the  change.  The  roasting  takes  place  in  four 
stages  as  shown  below. 

During  the  first  stage,  of  one  and  one-half  hours,  the  draft  is 
checked,  the  side  doors  kept  open,  and  the  charge  held  at  a  low 
temperature.  The  charge  becomes  evenly  heated  throughout,  and 
glows  from  the  oxidation  of  Cu2S  to  Cu20. 

During  the  second  stage,  of  one  and  one-half  hours,  the  heat 
is  increased  and  the  charge  constantly  rabbled.  Iron  sulphate  is 
decomposed  with  the  consequent  formation  of  copper  sulphate.  The 
charge  swells  and  becomes  spongy  by  the  formation  of  this  salt. 

In  the  third  stage,  of  a  like  duration,  the  temperature  is  increased 
for  an  hour  until  tests   show  that  the  silver  is  'out,'  that  is,   in 
the  form  of  sulphate.     The  following  reaction  occurs. 
CuSO4  +  Ag2O  =  Ag2SO4  +  CuO 

During  the  fourth  stage  the  temperature  is  kept  constant.  The 
charge  is  gathered  and  pressed  down  with  a  heavy,  long-handled, 
iron  paddle  to  break  the  lumps,  and  then  vigorously  stirred  to 


252  THE    METALLURGY 

oxidize  the  remaining  Cu20  to  CuO,  and  decompose  copper  sulphate. 
The  temperature  is  not  further  increased,  since  it  would  decompose 
silver  sulphate  forming  silver  oxide  rendering  the  silver  again 
insoluble. 

The  progress  of  the  roast  is  tested  by  dropping  small  samples 
from  time  to  time  into  hot  water.  Soluble  sulphates  dissolve  in 
the  hot  water;  and  in  the  tests  made  early  the  solution  becomes 
deep  blue.  Later,  as  the  silver  sulphate  begins  to  form,  it  is 
immediately  reduced  to  silver  spangles  by  the  cuprous  oxide  present. 
As  the  roasting  advances  during  this  stage,  the  copper  sulphate 
decomposes,  and  the  solution  becomes  less  blue  in  the  test  and  the 
silver  spangles  increase  and  afterward  diminish.  During  the  fourth 
stage  the  Cu20  is  changed  to  CuO  and  the  spangles  no  longer  show. 
A  light  blue  color  of  the  solution  remains,  due  to  the  presence  of 
a  little  copper  sulphate,  which  indicates  that  the  silver  sulphate 
is  not  itself  becoming  decomposed.  A  sample  thus  roasted,  showed 
by  analysis  2.5%  FeS04  and  ZnS04;  0.6%  CuS04,  and  1.73%  Ag2SO4 
(348  oz.  silver  per  ton),  so  that  there  was  left  in  the  matte  (there 
being  no  loss  of  weight  in  roasting  matte)  52  oz.  per  ton  or  13% 
of  the  silver  in  insoluble  form. 

Leaching. — This  is  performed  in  tanks  having  filter  bottoms  and 
holding  1000  Ib.  of  matte.  The  roasted  matte  charged  into  the 
tanks  /'  is  leached  with  hot  water  to  dissolve  the  sulphate  above 
described.  The  filtrate  goes  to  a  series  of  boxes,  h,  containing  copper 
plates  upon  which  the  silver  precipitates  in  the  form  of  white 
shining  crystals.  The  silver-free  solution,  containing  in  addition 
to  the  original  copper  sulphate  that  which  it  has  taken  from  the 
copper  plates,  goes  to  tanks  i  where  the  copper  is  precipitated  upon 
scrap-iron  employed  to  recover  the  copper.  The  final  solution  is 
rejected. 

The  cement-silver  from  the  precipitating  boxes  is  transferred 
to  a  tank  and  dilute  sulphuric  acid  is  added.  It  is  boiled  by  forcing 
in  a  mixture  of  air  and  steam  from  an  injector.  The  treatment 
oxidizes  and  dissolves  the  traces  of  copper  still  retained  by  the 
silver  crystals,  and  keeps  them  in  agitation  at  the  boiling 
temperature  of  the  acid  mixture.  The  copper  sulphate  solution  is 
run  off  and  the  residue  repeatedly  washed  by  decantation  with 
hot  water  to  free  it  entirely  from  copper.  It  is  transferred  to  a 
long  pan  over  a  coal  fire  for  drying  and  is  then  melted  down  in 
crucibles  in  a  wind-furnace  and  obtained  in  ingots  999  to  999.5 
fine. 

Residue  from  the  leaching  tanks. — The  extracted  matte  remaining 


OF    THE    COMMON    METALS.  253 

in  the  tanks  /',  still  retaining  52  oz.  silver  per  ton  as  above  stated, 
freed  from  sulphates,  and  composed  mainly  of  iron  and  copper 
oxides,  is  sent  to  a  reverberatory  furnace  k,  to  form  copper  matte. 
The  slag  produced  in  the  treatment  goes  back  to  the  ore-smelting 
furnace  a,  while  the  matte  tapped  into  sand  molds,  is  sent  to  the 
reverberatory  furnace  I  to  be  treated  by  the  English  process  of 
making  'best-selected  copper.'  Here  the  matte,  in  large  lumps,  is 
piled  up  in  the  furnace  near  the  bridge,  and  exposed  to  a  flame 
made  oxidizing  by  an  excess  of  air  admitted  through  the  fire  and 
through  openings  in  the  bridge  and  roof  of  the  furnace.  The  effect 
is  to  'roast'  the  matte  as  the  lumps  slowly  melt  and  the  drops  of 
liquefying  matte  come  in  contact  with  the  air.  Finally  the  whole 
charge  becomes  melted,  and  the  copper  oxide  which  has  been 
formed,  acts  on  the  unoxidized  copper  sulphide  of  the  matte  as 
follows : 

2Cu2O  +  Cu,S  =  6Cu-+  S02 

The  aim  is  to  extend  the  roasting  only  so  far  as  to  obtain  in 
the  form  of  metallic  copper  one-fifteenth  of  the  total  matte.  When 
the  charge  is  tapped  from  the  furnace  into  the  sand  molds,  the 
copper  is  found  in  the  form  of  plates  or  bottoms  in  the  first  of 
the  molds  beneath  the  lighter  matte.  The  bottoms  absorb  the 
impurities  such  as  arsenic,  antimony,  lead,  and  bismuth,  practically 
all  the  gold  (100  to  200  oz.  per  ton),  and  some  of  the  silver.  On 
the  other  hand,  the  matte  has  risen  to  the  grade  of  white-metal 
of  75%  copper,  and  carries  90  to  100  oz.  silver  but  not  more  than 
0.2  oz.  gold  per  ton. 

To  prepare  it  for  the  extraction  of  the  silver,  the  matte  is  again 
given  a  sulphatizing  roast,  but  in  a  different  furnace  from  the  one 
used  for  the  first  matte.  The  residue  after  this  second  treatment, 
principally  a  copper  oxide  containing  10  oz.  silver  per  ton  is 
sold  to  the  oil  refiners.  At  the  Argo  works  the  bottoms,  formerly 
treated  by  a  secret  process  for  the  extraction  of  the  precious 
metals,  is  now  more  satisfactorily  treated  by  the  processes  of 
electrolytic  refining.  (See  'The  Electrolytic  Refining  of  Copper'.) 

64.     THE    HYPOSULPHITE    LIXIVIATION    OF    SILVER    ORE. 
(PATERA  PROCESS.) 

Hyposulphite  lixiviation  can  be  practised  upon  ore  containing 
simple  or  compound  sulphides  of  silver  that  have  undergone  a 
preliminary  chloridizing  roasting.  The  silver  sulphides,  in  the 
roasting,  become  converted  into  silver  chloride.  The  process  also 
applies  to  silver  ore  already  containing  the  silver  as  chloride.  Free- 


254 


THE    METALLURGY 


milling  ore,  such  as  oxidized  ore  containing  the  silver  in  the  native 
state,  or  as  chloride,  or  to  some  extent  as  argentite,  are  preferably 
treated  by  milling  and  amalgamation.  Native  silver  and  silver 
sulphide  in  a  favorable  form  can  be  recovered  by  milling  and 
-amalgamation,  whereas  by  hyposulphite  extraction,  they  would 
remain  insoluble.  The  process  is  not  suited  to  the  treatment  of 
gold  ore.  The  extraction  of  gold  is  slow;  usually  less  than  50  to 
70%.  One  of  the  most  useful  applications  is  to  the  treatment  of 


40 


Fig.    114. 


Section  onB-3 

PLAN  AND   ELEVATION   OF  LIXIVIATION   PLANT   FOR 
SILVER  ORES. 


argentiferous  blende  that  has  been  hand-picked  or  concentrated 
from  galena,  and  may  still  contain  lead  up  to  eight  per  cent. 

The  hyposulphite  process  is  based  upon  the  fact  that  silver 
chloride  readily  dissolves  in  dilute  solutions  of  sodium  hyposulphite. 
The  chloridizing  roasting  is  unquestionably  the  most  important  part 
of  the  process,  and  the  chief  attention  and  study  is  to  be  given  it. 

The  process  consists  in  crushing  the  ore,  roasting  it,  and  treat- 
ing the  roasted  ore  in  filter-bottom  vats,  first  with  water  to  remove 
the  soluble  chlorides  and  sulphates  of  the  heavy  metals  (base-metal 


OF    THE    COMMON    METALS. 


255 


leaching),  then  leaching  with  a  dilute  solution  of  sodium  hypo- 
sulphite to  dissolve  the  silver  chloride.  Silver  sulphide  is  pre- 
cipitated from  the  filtrate  with  sodium  sulphide,  dried,  and  roasted 
to  remove  the  sulphur,  and  the  residue  is  sent  to  the  smelting  works, 
or  treated  in  an  English  cupelling  furnace. 

The  treatment  of  the  ore  to  the  point  where  it  is  stored  on  the 
cooling  floor,  has  been  described  under  dry  silver-milling,  which 
see. 

Fig.  114  and  115  show  the  plan  and  elevation  of  a  plant  for 
the  treatment  of  roasted  ore  by  hyposulphite  lixiviation.  At  the 
high  level  there  are  six  leaching  tanks,  each  20  ft.  diam.  by  6  ft. 


Fig.    115.      SECTIONAL   ELEVATION  OF   PLANT   FOR   THE  LIXIVIATION 

OF   SILVER  ORES. 

deep  and  each  holding  60  tons.  These  are  commanded  by  a  double 
track  over  which  the  ore  is  brought  to  them.  At  the  next  lower 
level  are  the  six  precipitating  tanks  13  ft.  diam.  by  10  ft.  deep,  and 
at  a  still  lower  level  are  four  sump-tanks,  serving  also  as  storage 
tanks,  whence  the  barren  solution  is  returned  by  centrifugal  pump 
to  be  again  used.  The  tanks  rest  upon  trestles,  so  as  to  be  accessible 
for  inspection  and  repair.  The  precipitate  is  treated  in  a  building 
not  shown. 

Base-metal  leaching. — The  cooled  roasted  ore  is  sprinkled  with 
water  from  a  hose,  and  piled  at  the  side  of  the  track  awaiting  to 
be  trammed  to  the  leaching  vats.  Fig.  116  represents,  in  plan  and 
sectional  elevation,  a  leaching-vat  having  a  false-bottom  of  wooden 
strips  covered  with  cocoa  matting  and  canvas.  Into  the  vat  is 
dumped  the  ore,  which  is  leveled  a  foot  below  the  top  of  the  tank. 

If  the  wash-water  were  admitted  above  and  allowed  to  leach 
downward,  it  would  dissolve  the  silver  chloride,  carrying  it  away 


256 


THE    METALLURGY 


in  the  wash-water.  To  avoid  this,  water  is  admitted  below  the 
false-bottom  and  made  to  ascend  through  the  charge  until  the  vat 
is  full.  The  wash-water  is  then  allowed  to  flow  out  at  the  bottom 


Fig.    116.      VAT   FOR  HYPOSULPHITE   LIXIVIATION. 

of  the  vat,  and  fresh  water  is  admitted  on  top.  By  this  means 
concentration  of  the  silver  in  the  solution  at  the  lower  part  of 
the  vat  is  avoided,  and  that  in  the  upper  part  is  soon  counteracted 
by  the  dilution.  The  precaution  does  not  prevent  the  dissolution 


OF    THE    COMMON    METALS.  257 

of  a  small  amount  of  silver  chloride  in  the  salty  water,  and  the 
first  of  the  discharge  is  run  to  a  special  precipitation-vat  at  the 
end  of  the  row.  The  flow  is  through  the  outlet  d,  Fig.  116,  and  the 
hose  ?/,  opened  for  the  purpose,  and  thence  by  launder  g  (marked 
d  in  the  plan  Fig.  114)  to  the  vat.  Here  the  silver  is  precipitated 
by  sodium  sulphide. 

When  the  quantity  of  copper  in  the  ore  warrants,  the  clear 
solution,  decanted  from  the  silver  precipitate,  is  passed  through 
launders  containing  scrap-iron  to  recover  the  copper.  In  any  case, 
the  washing  of  the  ore  in  the  leaching  vat  continues  two  or  three 
days,  the  hose  being  turned  into  the  next  compartment  of  the 
launder  g  (e  of  Fig.  114)  and  the  last  of  the  wash-water  run  to 
waste.  The  completion  of  the  washing  is  so  thorough  as  to  entirely 
remove  the  soluble  base-metal  salts. 

Silver  leaching. — A  cold  dilute  solution  of  sodium  hyposulphite 
('hypo  solution')  is  admitted  to  the  leaching  vat  above  the  ore.  It 
enters  from  the  main  distributing  launder  c,  by  a  short  rubber-hose, 
the  end  of  which  can  be  raised  to  stop  the  flow.  The  leaching 
proceeds  rapidly,  the  outlet  d  being  kept  open. 

The  hypo-solution  dissolves  silver  chloride,  reacting  as  follows: 

2Na2S2O3  +  2AgCl  =  AgS203Na2S2Os  +  2NaCl 

Not  only  the  silver  chloride,  but  also  silver,  silver  oxide,  silver 
arsenate,  silver  antimonate,  and  gold  pass  into  solution.  Copper 
chloride  dissolves  much  like  silver.  Lead  sulphate  and  calcium 
sulphate  dissolve,  but  the  solvent  power  of  the  hypo-solution  is 
diminished  by  the  presence  of  lead  sulphate  or  sodium  sulphate, 
and  particularly  of  caustic  alkalis  and  alkaline  earths  such  as 
quicklime,  presence 'of  the  latter  being  due  to  the  roasting  of  the 
limestone  in  the  ore. 

The  lower  part  of  the  ore  in  the  roasting  vat,  after  washing, 
contains  15  to  20%  moisture,  causing  the  first  of  the  hypo-solution 
to  be  greatly  diluted.  To  avoid  weakening  the  whole  stock  of 
solvent  the  portion  first  traversing  the  ore  is  rejected.  As  soon 
as  the  effluent,  tested  with  sodium  sulphide,  shows  traces  of 
precipitation,  the  hose  is  transferred  to  launder  g  and  the  solution 
admitted  to  the  special  precipitation-vat  and  treated  as  already 
described.  When  samples  from  the  discharge  yield  distinct 
precipitate,  the  liquid  is  run  to  one  of  the  regular  precipitating  vats. 

Leaching  proceeds  until  tests  show  that  the  silver  is  extracted. 
The  flow  of  hypo  is  then  stopped,  and  a  water-wash  run  on.  This 
removes  the  hypo-solution,  and  by  changing  the  hose,  the  flow  can 
be  directed  to  one  of  the  four  sumps,  or  stock-tanks,  at  the  lowest 


258  THE    METALLURGY 

level,  until  the  solution  becomes  weak  and  does  not  pay  to  save. 
To  avoid  excessive  solvent,  and  to  provide  adequate  circulation 
the  outlet  d,  Fig.  116,  of  the  tank  is  connected  through  t  with  the 
Koerting  injector  k,  which  produces  a  partial  vacuum  below  the 
filter-bottom,  and  forces  the  solution  into  the  launder  g  (c  in  Fig. 
114),  whence  it  is  admitted  again  to  the  charge.  The  final  draining 
is  through  the  hose  u  to  the  launder  g.  When  lead  or  copper  is 
contained  in  the  ore  it  is  advisable  to  use  dilute  hypo-solution  (1.0 
to  1.2%)  circulating  in  large  volume.  The  dilute  solution  acts  upon 
the  silver  salts  in  preference  to  the  lead  salts  (selective  affinity)  and 
the  former  can  be  extracted  without  serious  dissolution  of  the  latter. 
The  greater  part  of  the  solution  is  extracted  during  the  first  8  to 
12  hours  of  the  silver  leaching,  and  is  known  as  the  sweet  solution, 
because  of  the  sweet  taste  of  silver  chloride  dissolved  in  the  hypo- 
sulphite. The  metal  precipitated  from  this  solution  will  be  80% 
silver ;  on  the  other  hand,  the  precipitate  from  solution  flowing  from 
base  ore,  during  the  remainder  of  the  period,  is  poor  in  silver, 
hence  the  advantage  of  using  dilute  solutions,  and  of  leaching 
rapidly.  Under  these  conditions,  with  ores  containing  much  lead 
and  copper,  we  may  expect  to  obtain  a  solution  that  will  yield 
1.5  to  2.4  oz.  silver  per  ton  when  30-oz.  ore  has  been  successfully 
chloridized. 

Precipitation. — A  strong  solution  of  sodium  sulphide  is  added  to 
the  silver-bearing  solution  in  the  precipitation  tank  and  well  stirred 
by  hand  or  mechanically.  The  quantity  of  sodium  sulphide  should 
be  such  as  to  leave  a  slight  excess  of  silver  solution  rather  than 
an  excess  of  sodium  sulphide.  The  proportion  is  determined  by 
testing  a  sample  with  a  drop  of  sodium  sulphide. 

The  silver  precipitates  as  sulphide  as  indicated  in  the  following 
reaction : 

Ag2S203  Na2S203  +  Na2S  =  Ag2S  +  2Na2S208 

Sodium  hyposulphite  is  regenerated :  Gold,  copper,  zinc,  and  lead 
are  thrown  down. 

When  the  proper  time  has  arrived,  the  precipitate  is  allowed 
to  settle,  and  the  supernatant  solution,  drawn  down  to  1%  to  2 
ft.  above  the  bottom  of  the  precipitating  tank,  is  run  to  one  of 
the  sump  or  stock-tanks  below,  making  use  of  the  floating  hose, 
see  Fig.  76. 

The  effect  of  adding  the  sulphide  is  to  regenerate  the 
hyposulphite,  making  it  again  suitable  for  use.  If  necessary,  it 
may  be  further  strengthened  by  the  addition  of  the  dry  salt.  Sodium 
sulphide  solution  may  be  made  at  the  works  by  dissolving  caustic 


OF    THE    COMMON    METALS.  259 

soda  in  its  own  weight  of  water  at  80°  C.  in  a  3-ft.  diam.  iron 
kettle,  adding  gradually  powdered  sulphur  to  the  solution.  The 
addition  of  sulphur  causes  the  liquid  to  increase  to  two  or  three 
times  the  original  bulk,  and  in  the  beginning  the  pot  should  be 
only  one  quarter  full.  The  sulphur,  60%  the  weight  of  the  caustic 
soda,  dissolves  in  a  few  minutes.  The  sodium  sulphide  solution 
resulting  is  poured  into  a  mold  in  which  it  solidifies.  For  use  it  is 
dissolved  in  water.  The  sulphur  needed  varies  from  4  to  9  Ib.  per 
ton  of  ore  treated,  and  the  sodium  hyposulphite  2  to  4  Ib.  per  ton. 

Treatment  of  the  precipitate. — The  residue  in  the  precipitating 
tank  is  discharged  through  the  pipe  /,  Fig.  114  and  115,  to  a  collect- 
ing tank  in  a  building  not  shown.  From  the  collecting  tank  it  is 
pumped  to  a  filter  press,  and  after  pressing,  the  damp  precipitate  is 
put  in  a  reverberatory  furnace  16  by  6l/2  ft.  hearth  dimensions  hav- 
ing a  small  grate  to  give  a  moderate  heat.  It  is  heated  gradually  in 
the  furnace  until  quite  dry,  after  which  the  temperature  is  increased 
to  burn  the  sulphur. 

The  dried  precipitate  is  14  to  35%  silver  and  15  to  27%  copper. 
Where  ore  is  comparatively  free  from  copper  and  lead,  the  sil- 
ver may  rise  to  45  or  55%  in  the  precipitate.  In  the  following 
table  we  give  an  analysis  of  sulphides  from  the  Marsac  mill,  Park 
City,  Utah. 

Per  cent. 

Ag   (10,124  oz.  per  ton) 34.78 

Au   (11.7  oz.  per  ton) 0.04 

Cu    21.60 

Pb 0.50 

Fe    0.75 

Sb 0.18 

A1203     0.25 

Si02     0.25 

S    20.74 

The  dried  and  roasted  sulphides  are  generally  sold,  and  go  to 
smelting  works  where  they  are  smelted  directly,  being  fed  into  the 
blast-furnace  in  sacks,  or  treated  in  an  English  cupelling-furnace 
as  follows : 

Thrown  upon  the  molten  lead-bath  in  the  hot  furnace,  a  few 
shovelfuls  at  a  time,  the  roasted  precipitate  melts.  The  silver  enters 
the  lead.  The  base  metals,  undergoing  a  scorifying  action,  enter 
the  litharge.  The  slag  continuously  forming  flows  from  the  furnace, 
and  being  a  litharge  slag,  and  containing  silver,  is  sent  to  the  silver- 


260  THE    METALLURGY 

lead  blast-furnace.  When  all  the  sulphide  has  been  treated,  the 
lead  of  the  molten  bath  is  cupelled.  The  silver  is  left  behind  and 
is  tapped  from  the  furnace  into  molds  to  form  bars  or  ingots. 

Cost  of  treatment. — At  Sombrerete  the  ore  contains  9  to  10% 
lead  as  galena,  blende,  chalcopyrite,  silicious  gangue,  and  silver 
sulphide.  The  ore  assays  41.9  oz.  silver  per  ton.  It  loses  by 
roasting  4.8%  of  this.  The  extraction,  figured  on  the  raw  ore,  is 
82.5%.  The  treatment  cost  is  as  follows: 

Crushing    $1.36 

Roasting  (including  the  use  of  6%  salt) ....  2.68 

Labor  at  the  leaching  plant 0.27 

Chemicals     0.30 

Superintendence     1.02 

Heating,  lighting,  pumping,  and  repairs.  . .  .  0.08 


Cost  per  ton $5.71 

Clemes  has  given  the  cost  in  Mexico,  at  works  treating  40  to  50 

tons  per  day,  as  being  1*8  to  W  in  the  coast  districts,  and  f*ll  to  W2 

in  the  mountain  districts.     In  small  works  the  ratio   of  the  cost 

of  superintendence  to  the  total  cost  is  high.     The  small  works  are 

r\V      usually  directed  by  the  owners. 

W  65.     THE  RUSSELL  PROCESS. 

This  is  a  modification  of  the  Patera  process,  principally  by  the 
use  of  another  solution,  in  addition  to  the  hypo-solution,  for  the 
extraction  of  the  silver.  By  mixing  in  solution  two  parts  of  the 
hypo-salt  with  one  of  copper  sulphate  we  obtain  a  double  salt 
Na2S2O33Cu203,  called  the  extra-solution.  It  has  a  solvent  power 
nine  times  asAgreat  as  that  of  the  ordinary  hypo-solution  for  native 
silver,  silver  sulphides,  silver  arsenides,  and  silver  antimonides. 
In  the  case  of  an  imperfectly  roasted  ore  the  use  of  the  extra- 
solution  insures  the  extraction  of  more  silver  from  the  compounds 
mentioned  than  could  be  obtained  by  the  use  of  ordinary  hypo- 
solution. 

The  practice  at  the  Marsac  mill,  Park  City,  Utah  (See  Fig.  117), 
is  as  follows:  The  silicious  ore,  containing  5.3%  copper,  0.39% 
lead,  37.3  oz.  silver,  and  0.05  oz.  gold  per  ton,  is  dried  and  dry- 
milled,  using  a  30-mesh  screen,  and  obtaining  an  output  of  2% 
tons  per  stamp  in  24  hours.  It  is  then  mixed  with  8.9%  salt  and 
roasted  in  a  chloridizing  furnace,  then  delivered  to  the  point  marked 
'cooling  floor',  Fig.  117. 


OF    THE    COMMON    METALS. 


261 


The  ore,  when  cooled  and  chloridized,  is  charged  to  the  ore- 
lixiviation  vat  17  ft.  diam.  by  9  ft.  deep,  holding  72  tons:  It  is 
here  treated  by  a  water-wash  for  19  hours,  using  0.56  ton  water 
per  ton  ore,  the  water  percolating  the  charge  at  the  rate  of  4  in. 
per  hour.  Dissolving  the  base-metal  salts  and  a  part  of  the  silver 


BORDER  OF  OP  C*A  T/CW3  -» 
ist.  Preparation  of  the  Mater/mis 
xtraction 


4tt,.  Treatment  ef  ft-oefucrs 


Fig.    117.      FLOW-SHEET   OF   RUSSELL  PROCESS. 

chloride  the  percolating  wash-water  passes  to  the  silver  and  base- 
metal  precipitation  tank.  Here,  by  the  addition  of  sodium  sulphide, 
a  precipitate  is  obtained  containing  on  an  average  4.2%  lead,  3.9% 
copper,  and  1238  oz.  silver  per  ton.  This  is  separated  from  the 
contained  water  in  a  Johnson  filter-press,  removed  from  the  press, 
dried  and  shipped  to  smelting-works. 

The  ore  is  next  treated  with  a  1.5%  hyposulphite  stock-solution 
by  which  the  greater  part  of  the  silver  is  extracted.  The  first  of 
the  flow  goes  to  the  base-metal  precipitation  tank,  but,  as  soon  as 


262  THE    METALLURGY 

the  hypo-solution  begins  to  appear,  the  solution  is  diverted  to  the 
lead  precipitation  tank.  The  silver  leaching  lasts  81  hours. 

Now  the  extra-solution  containing  0.75%  bluestone  and  1.5% 
hypo  is  applied  to  the  ore,  and  further  silver  extracted  from 
undecomposed  sulphides  not  attacked  by  the  hypo-solution.  This 
operation  takes  27  hours,  the  filtrate  being  also  run  to  the  lead- 
precipitation  tank.  The  cycle  of  leaching  takes  130  to  150  hours. 

In  the  lead  precipitation  tank  lead  carbonate,  which  is  insoluble 
in  hypo-solution,  is  precipitated  by  the  addition  of  5  Ib.  sodium 
carbonate  per  ton  of  solution,  precipitation  occuring  in  result  of 
the  following  reaction : 

Na2C03  +  PbCl2  =  2NaCl  +  PbC03 

The  precipitate  contains,  on  the  average,  32%  lead  and  526  oz.  silver 
per  ton.  After  the  settling,  the  supernatant  solution  is  transferred 
to  the  tank  marked  'silver-gold-copper-precipitation,'  while  the  wet 
precipitate  is  filter-pressed,  dried,  and  sold  to  be  smelted. 

The  solution  in  the  silver-precipitating  tank  is  now  treated  as 
in  the  Patera  process  with  just  enough  sodium  sulphide  to  bring 
down  the  precipitate  of  silver  sulphide  containing  35%  silver,  20% 
copper,  and  20%  sulphur.  The  clear  solution  is  pumped  back  to  the 
hyposulphite  stock-solution  tank  to  be  again  used,  while  the 
precipitate  is  pumped  to  the  filter-press,  the  solution  removed,  and 
the  resultant  moist  precipitate  dried  and  roasted  at  a  low  heat  to 
remove  sulphur. 

Thus  the  filter-press  and  dryer  treat  the  base-metal  precipitate, 
the  precipitated  lead  carbonate,  and  the  silver  precipitate,  one 
after  another.  The  products  are  sacked,  being  sold  to  smelters,  who 
either  smelt  them  directly  or  treat  them  in  an  English  cupelling 
furnace  as  described  in  the  Patera  process. 

The  extraction  by  the  Kussell  process  is  as  follows: 

Per  cent. 

In  base-metal  sulphides 6.2 

In  lead   carbonates 2.6 

In  silver  precipitates 75.7 

In  sweepings  of  the  mill.  . 0.5 

Total   recovery 85.0 

The  Russell  process  has  proved  successful  in  exceptional  cases 
only.  At  Sombrerete  and  at  Cusihuiriachic,  Mexico,  it  has  been 
displaced  by  the  Patera  process,  while  failures  have  occured  else- 
where. Compared  with  the  Patera  process,  the  cost  of  chemicals 
is  greater  (92c.  against  42c.),  the  plant  is  more  complicated,  and 


OF    THE    COMMON    METALS.  263 

greater  skill  is  needed  to  work  it  successfully.  It  can  be  applied  to 
oxidized  ores,  or  to  those  that  have  been  subjected  to  an  oxidizing 
roast,  and  though  the  yield  of  silver  is  greater  where  the  extra 
solution  is  used,  yet  this  is  offset  by  the  consumption  of  copper 
sulphate.  In  presence  of  much  galena  or  blende  the  extra-solution 
extracts  silver  no  better  than  the  ordinary  hypo-solution.  In 
presence  of  much  lime,  silver  is  but  slowly  extracted  and  copper 
sulphate  is  consumed. 

Cost. — The  cost  by  the  process,  based  upon  an  output  of  100 
tons  daily,  is  as  follows: 

Crushing  and  roasting $4.62 

Labor  in  leaching 0.83 

Tools,  lighting,  pumping,  and  heating 0.12 

Chemicals     0.92 

Kepairs  and  superintendence 0.18 

Assaying    0.08 

Treatment  of  products 0.10 


Total     $6.85 

66.     CYANIDATION  OF  SILVER  ORES 

Silver  ores  carrying  gold,  and  in  which  the  silver  occurs  as 
chloride,  argentite,  or  stephanite,  have  been  sucessfully  cyanided. 
Thus  at  Chloride  Point,  Utah,  where  silver  occurs  as  the  chloride, 
the  extraction  is  71%  ;  at  the  Palmarejo  mine,  Chihuahua,  Mexico, 
the  extraction  is  54%  silver  and  96%  gold;  at  El  Salvador, 
85  to  90%  silver  and  90  to  92%  gold;  while  at  Guanajuato, 
a  silver  ore,  containing  silver  sulphide  and  a  little  gold,  is  treated 
with  an  extraction  of  87%  of  the  total  metal  of  value. 

Of  the  silver  minerals,  native  silver,  in  particles  so  large  as 
to  be  visible,  is  insoluble  in  potassium  cyanide  in  any  reasonable 
time.  Silver  chloride,  bromide,  and  argentite,  are  readily  soluble. 
Ruby  silver,  stephanite,  and  frieslebenite  are  sparingly  soluble  in 
potassium  cyanide  but  readily  soluble  in  murcurous  potassium 
cyanide  solution. 

Important  matters   in   cyaniding   silver  ore   are  the  following: 

(a)  A  long  time  (10  to  25  days)  in  the  case  of  sand  treatment 
is  needed  for  leaching.  For  slime  treatment  96  hours  would 
suffice  for  a  complete  cycle  in  which  time  a  higher  percentage  of 
extraction  would  be  obtained  than  by  a  14-day  treatment  of  the 


264  THE    METALLURGY 

corresponding   sand.      The   silver   compounds    are    more    difficultly 
soluble  than  gold,  and  a  larger  amount  must  be  dissolved. 

(b)  Thorough   oxygenation  is  necessary,  not  only  because   of 
the  large  amount  of  silver  present,  but  because  silver  compounds 
need  at  least  initial  oxidation  to  become  properly  soluble  in  cyanide 
solution.     Hence  an  advantage  is  secured  by  the  double  treatment 
of  sand,  as  described  under  the  'second  method'  for  the  treatment 
of  gold  ore.     Also  during  leaching,  if  the  solution  be  allowed  to 
sink   several  inches  below  the  top   of  the   charge,   before   another 
wash-solution  is  run  on,  air  is  drawn  and  penetrates  the  ore,  and 
the  solution  following  forces  the  air  downward  through  the  ore. 
In  the  treatment  of  the  slime  the  pulp  may  receive  through  aeration 
by  agitating  with  air. 

(c)  Stronger  solution  is  used  than  for  the  treatment  of  gold 
ore.     Thus  the  first  or  strong  solution  may  be  0.1%,  the  weak  one 
0.25%,  while  for  gold  ore,  a  0.5%  solution  would  be  called  strong 
and  0.05%  weak. 

(d)  The  consumption  of  potassium  cyanide  is  higher  than  in 
the  treatment  of  gold  ores.     It  varies  from  2.5  to  4  Ib.  per  ton  as 
compared  with  0.4  to  0.8  Ib.   consumed  in  the  treatment  of  gold 
ores. 

(e)  The  precipitation  of  silver  from  cyanide  solution  by  zinc 
shaving  presents  no  difficulties,  and  is  practically  complete.    Despite 
the  fact  that  a  relatively  great  amount  of  silver  has  to  be  precipated, 
as  compared  with  gold,  no  more  zinc  is  consumed. 

The  separation  and  leaching  of  the  sand  should  be  discontinued 
in  many  cases,  since  leaching  by  cyanide  solution  only  slowly 
dissolves  the  fine  mineral  particles  that  are  enveloped  in  the  quartz 
grains,  and  the  time  that  economically  can  be  allowed  for  extraction 
is  inadequate.  When  the  ore  is  tough  and  hard  and  the  added 
expense  of  fine  grinding  offsets  the  additional  extraction,  the  above 
practice  would  not  apply.  In  most  ores  a  part  of  the  silver  sulphide 
is  so  finely  disseminated  in  the  gangue  that  it  becomes  necessary 
to  grind  the  ore  to  pass  a  150-mesh  screen  before  the  sulphides 
are  liberated  sufficiently  from  the  enclosing  quartz  to  permit  a 
satisfactory  extraction  of  the  silver  by  cyanidatioii. 

While  mechanical  agitation  of  slime  in  tanks  provided  with 
stirring-arms  is  common,  agitation  can  be  more  economically  and 
effectively  executed  by  the  use  of  Brown 's  agitating-tank,  Fig  118. 
This  is  provided  with  an  air-lift  consisting  of  a  12-in.  pipe  extending 
nearly  to  the  bottom  of  a  tall  conical-bottom  tank  13  ft.  diam.  by 
55  ft.  high.  A  central  pipe  c  supplies  air  under  pressure.  The  air 


OF    THE    COMMON    METALS. 


265 


bubbling  through  the  pulp  causes  a  flow  upward,  and  creates  a 
circulation.  By  this  method  a  large  proportion  of  fine  sand  can 
be  suspended  in  the  slime  pulp  as  would  not  be  possible  by  the 


7;cN  OF  BROWN  AGITATOR 

Fig.    118.      SECTIONAL   ELEVATION  OF  BROWN  AGITATOR. 

usual    stirring    and    pumping    method    of    agitation.      Upon    the 
dissolution  of  the  gold,  the  slime  is  filter-pressed. 

The  tendency  in  advanced  practice  is  to  classify  and  finely  grind 
the  sand  to  pass  a  150-mesh  screen,  to  add  this  to  the  slime,  and 


266  THE    METALLURGY 

to  agitate  and  filter-press  the  united  product.     The  fine  sand  thus 
present  with  the  slime  makes  it  more  easily  filtered. 

67.     CYANIDATION  OF  SILVER  ORES  AT  GUANAJUATO. 

The  ore  consists  of  pyrite  associated  with  a  quartz  gangue  with, 
disseminated,  the  silver  sulphides  argentite,  stephanite,  and 
pyrargyrite.  It  carries  14  oz.  silver  and  0.08  oz.  gold  per  ton.  The 
treatment  is  that  described  in  outline  in  the  'second  method'  of 
cyaniding  gold  ores,  except  that  here  the  ore  is  crushed  in  barren 
cyanide  solution. 

The  ore  is  fed  from  storage  bins  to  Blake  breakers  and  is  crushed 
to  2-in.  size.  It  is  elevated  by  a  bucket  elevator  to  the  sorting-room, 
and  passed  over  a  grizzly  to  separate  the  fine  while  the  lump-ore 
falls  upon  a  sorting  belt.  From  10,000  tons  of  ore,  the  monthly 
output,  ore  sorters  pick  out  the  waste,  leaving  8000  tons  which 
is  hauled  in  cars  drawn  by  industrial  locomotives  to  the  storage- 
bins  of  the  stamp-mill,  the  total  cost  being  8c.  per  ton.  Thus  the 
grade  of  the  ore  is  increased  at  the  start  and  the  quantity  for 
treatment  diminished. 

As  the  ore  (260  tons  daily)  is  dumped  into  the  storage-bins, 
\%  quicklime  is  added  to  neutralize  the  free  sulphuric  acid  in  the 
ore  and  furnish  protective  alkalinity  to  the  cyanide  solution.  The 
ore  is  fed  automatically  to  the  stamps,  where  7.2  tons  of  barren 
solution  (0.025%  cyanide)  per  ton  of  ore  is  used.  Thus  the  silver 
begins  to  dissolve  even  while  the  ore  is  being  crushed.  From  the 
battery  the  ore  passes  through  a  50-mesh  screen  at  the  rate  of  3.3 
tons  per  stamp  daily.  The  smaller  tonnage  compared  with  the 
output  elsewhere  is  due  to  the  fine  screen  used.  No  portion  of  this 
pulp  is  subsequently  re-ground  and  about  60%  will  pass  200-mesh. 
The  rich  parts  of  the  ore  are  the  more  friable,  and  consequently 
enter  the  finer  sizes. 

The  pulp  from  each  five-stamp  mortar  flows  through  spitzkasten 
each  having  two  compartments  22  in.  square  by  22  in.  deep.  The 
overflow  of  slime  goes  to  the  slime-tanks,  while  the  spigot-discharge, 
or  underflow,  unites  and  passes  to  the  Wilfley  concentrating  table. 
The  concentrate  from  the  table  goes  to  storage-tanks  in  the 
concentrate  room,  the  slime-overflow  from  the  last  five  feet  of  the 
"Wilfley  table  goes  to  the  slime-tanks,  and  the  remainder  of  the  tailing 
to  a  Frue  vanner  for  re-treatment.  The  vanner  concentrate  goes 
to  the  concentrate-room  and  is  mixed  with  the  Wilfley  concentrate. 
The  concentrate  amounts  to  2.5%  ore  and  contains  50%  total  value. 
It  carries  40%  Fe  and  6%  SiO2,  and  is  sent  to  the  smelter.  Mean- 


OF    THE    COMMON    METALS.  267 

while  the  vanner  tailing,  containing  all  the  sand,  flows  to  classifying 
cones  3l/2  ft.  diam.  provided  with  hydraulic  water.  The  overflow 
goes  to  the  slime-tanks. 

The  sandy  spigot-discharge,  containing  slime,  passes  by  a 
launder  to  the  sand-elevator  and  discharges  into  the  small  cone- 
separators  18  in.  diam.  provided  with  hydraulic  water.  Here  the 
overflow  removes  further  slime  which  goes  to  the  slime-vats.  The 
underflow,  or  spigot-discharge,  of  the  cone-separator  is  discharged 
into  the  collecting  vat  near  the  side.  Here  it  accumulates,  sloping 
away  gradually  to  the  opposite  side  of  the  tank.  A  grating,  such 
as  was  described  under  the  'second  method,'  and  shown  in  Fig.  86, 
permits  the  escape  of  slime-bearing  water.  The  grating  is  covered 
as  high  as  the  level  of  the  top  of  the  sand  with  a  canvas  roll- 
curtain  to  retain  the  sand.  Should  the  sand  still  contain  slime, 
a  boy  with  a  hose  can  wash  the  slime  from  the  surface  of  the 
depositing  sand  through  the  grating  to  the  slime-tanks.  The  sand- 
vats,  seven  in  number,  have  filter  bottoms  and  bottom-discharge 
valves  or  doors.  They  are  26  ft.  diam.  by  5  ft.  deep,  and  hold  90 
tons  of  classified  sand. 

Preliminary  treatment  of  sand. — The  collected  sand  receives  the 
first  treatment  in  the  vat  in  which  it  settles,  thus  making  the  vat 
one  of  preliminary  treatment.  To  fill  the  vat  requires  15  hours, 
and  to  drain  15  hours  more.  The  sand  retains  18%  moisture.  Upon 
the  charge  is  now  run  20  tons  of  solution  containing  0.3%  KCN  to 
which  has  been  added  20  grams  of  lead  acetate  per  ton.  This  is 
allowed  to  soak  12  hours.  While  adding  the  solution,  a  basket, 
containing  50  Ib.  dry  cyanide  salt,  is  placed  under  the  stream  and 
the  salt  slowly  dissolved  and  carried  into  the  charge.  The  addition 
is  sufficient  to  re-enforce  the  solution  to  the  strength  of  0.3%  KCN 
in  spite  of  the  dilution  by  the  moisture  in  the  charge.  Percolation 
and  drainage  now  follow.  Drainage  causes  aeration  of  the  charge. 
The  preliminary  treatment  is  completed  by  a  second  saturation, 
percolation  and  drainage  of  the  same  tonnage  and  strength  of 
solution  Of  the  first.  The  recovery  of  35  to  40%  of  the  original  value 
of  the  ore  is  the  result. 

Second  treatment  of  sand. — The  first  treatment  being  complete, 
four  discharging-doors  in  the  bottom  of  the  tank  are  opened,  and 
the  sand  containing  14%  moisture  is  shoveled  into  double  side- 
dumping  tram-cars  running  on  tracks  below  the  tanks,  and  above 
the  second  treatment  or  sand-tanks.  There  are  15  sand-tanks, 
each  26  ft.  diam.  by  6  ft.  deep,  being  made  more  capacious  to  provide 
for  the  sand  which  does  not  pack  in  them  as  in  the  collecting  tanks. 


268  THE    METALLURGY 

While  a  tank  is  being  charged,  1000  Ib.  quicklime  is  distributed 
uniformly  through  the  charge.  The  sand  is  next  saturated  with 
30  tons  of  strong  solution  to  which  is  added  50  Ib.  dry  cyanide,  and 
allowed  to  remain  in  the  tank  one  day.  The  outlet  valve  is  opened, 
and  the  solution  is  allowed  to  drain  through  the  charge.  The  sand 
is  saturated  a  second  time  with  the  regular  strong  working-solution 
of  the  same  strength  and  tonnage  as  before.  The  solution  valve 
is  opened  again  and  the  admission  of  strong  solution,  and  the 
percolation  simultaneously  proceed  6%  days.  The  sand  is  next 
washed  with  weak  solution  (0.15%  KCN)  two  days,  and  then  with 
water  18  hours.  The  two  bottom  doors  of  the  tank  are  opened 
and  the  sand  sluiced  out  and  carried  by  launder  to  the  dump.  The 
total  time  of  the  double  treatment  is  16  days,  during  which  9  Ib. 
lead  acetate  is  added  for  a  90-ton  charge.  The  progress  of  the 
double  treatment  of  a  charge  of  sand  is  shown  in  Fig.  119. 

Slime  treatment. — There  are  24  masonry  slime-tanks,  each  8  by 
9  ft.,  in  two  rows,  beneath  the  concentration-mill  floor.  The  stream 
of  slime,  separated  from  the  pulp  as  already  described,  flows  through 
one  row  of  the  tanks  until  filled.  The  stream  is  then  diverted  to 
the  other  row.  The  slime  in  the  first  row  is  permitted  to  settle 
quietly,  and  the  clear  supernatant  water  is  decanted  into  sump-tanks 
and  again  used  at  the  mill.  The  residual  slime  in  the  tank  contains 
three  parts  water  to  one  of  dry  slime,  and  amounts  to  one-half  the 
original  ore. 

Prom  the  settling  tanks  the  thickened  slime  goes  to  14  slime- 
treatment  tanks,  each  30  ft.  diam.  by  10  ft.  deep,  provided  with 
stirring  arms  for  mechanical  agitation.  The  charge  weighs  140  tons 
and  contains  35  tons  of  slime,  dry  weight.  As  the  stream  of  slime 
enters  the  tank  it  flows  upon  a  basket  containing  potassium  cyanide 
to  strengthen  the  solution  to  0.1%  KCN.  Then  200  Ib.  quicklime 
and  18  Ib.  lead  acetate  are  added.  The  lime  corrects  acidity,  gives 
protective  alkalinity,  and  assists  settling;  the  lead  acetate  decom- 
poses alkaline  sulphides  that  would  precipitate  silver  from  the 
cyanide  solution.  The  charge  fills  the  tank  to  a  depth  of  8  ft.,  and 
the  remaining  2  ft.  is  filled  with  0.1%  KCN  solution  from  the  stock- 
tanks.  The  solution  dilutes  the  slime  five  to  one,  a  proportion  found 
necessary  for  the  successful  treatment  of  the  charge. 

Agitation  now  proceeds  uninterruptedly  24  hours,  after  which  the 
charge  is  allowed  to  settle  six  hours.  By  this  time  three  feet  of  the 
clear  solution  can  be  decanted  to  the  adjacent  'clarifying'  or  clear- 
solution  tanks.  The  slime-treatment  tank  is  again  filled  with  0.1% 
KCN  solution,  adding  the  same  charge  of  lime  and  lead  acetate  as 


OF    THE    COMMON    METALS. 


269 


before,  and  agitation  is  resumed  and  continued  half  an  hour.  The 
pulp  is  allowed  to  settle ;  the  clear  solution  to  the  depth  of  18  in.  is 
decanted,  and  the  tank  filled  with  weak  solution  after  the  requisite 
addition  of  lime.  Six  washes  like  this  follow,  making  eight  in  all, 
followed  then  by  two  water-wTashes.  The  settled  slime  is  now 
pumped  to  one  or  two  settling  tanks  set  at  a  high  level,  where  it 
mixes  with  weak  solution  previously  pumped  into  the  tank  to  the 
depth  of  6  ft.  The  mixture  is  allowed  to  settle  as  long  a  time  as  the 
tank  can  be  spared  for  the  purpose.  The  clear  solution  is  decanted 
into  the  clarifying  tanks,  each  6  ft.  diam.  by  12  ft.  high,  with  the 
solution  decanted  from  the  slime-treatment  tanks.  The  pulp  remain- 


4  5%  Transfer?  *d  fo  First  Tf  9at/nent  Ksn  ra 


2  4  6  0   DAYS      10  /,?  14 

Fig.    119.      SAND-TREATMENT   CHART   AND   RECORD. 

ing  in  the  high-level  settling-tank  is  then  run  to  waste.  The  various 
decantations  entering  the  first  clarifying  tank,  flow  across  it  to  the 
second,  passing  through  three  curtain-screens  of  cocoa-matting  cov- 
ered with  burlap  which  form  partitions  through  which  the  solution  is 
strained. 

Treatment  of  the  silver  and  gold-bearing  solution. — The  strong 
solution  from  the  leaching  tanks  is  received  in  the  strong  gold- 
solution  tanks.  The  weak  solution  and  water-wash  of  these  tanks, 
as  well  as  the  solution  from  the  clarifying  tanks,  goes  to  the  weak 
gold-solution  tank,  and  after  passing  through  the  zinc  boxes  to  the 
strong-solution  and  weak-solution  sump-tanks  respectively.  From 
the  sumps  the  barren  solution  is  pumped  to  the  stock-tanks  to  be 


270  THE    METALLURGY 

re-enforced  to  working  strength  by  the  addition  of  dry  cyanide  and 
again  used.  Some  of  the  weaker  solution  (0.025%  KCN)  is  pumped 
to  the  reservoir  storage-tanks  for  use  in  the  batteries.  The  total 
weight  of  solution  used  is  15.5  tons  per  ton  of  ore,  or  for  the  daily 

tonnage  of  260  tons,  3(1  to  40  tons  of  solution  passes  through  the  zinc- 

i  o       °  Q 

boxes. 

The  clean-up. — The  zinc-boxes  are  cleaned  up  three  times  a  month 
as  described  under  the  first  method  of  cyaniding,  the  precipitate 
being  run  to  a  clean-up  sump-tank,  and  without  acid  treatment, 
pumped  through  a  filter-press.  The  cake  of  precipitate  from  the 
press  contains  30%  moisture,  52%  silver,  and  3%  gold.  The  15% 
remaining  consists  of  zinc,  lead,  copper,  sulphur,  and  silver.  The 
precipitate  is  mixed  with  fluxes  in  a  large  tray,  using  100  parts 
precipitate,  30  of  borax,  and  18  soda.  The  mixture  is  transferred  to 
shallow  sheet-iron  trays,  dried  by  steam,  then  charged  into  No.  300 
graphite  crucibles  and  melted.  The  molten  metal  is  cast  into  bars 
weighing  1000  oz.  each,  860  fine  in  silver  and  5  fine  in  gold,  the 
remaining  ^35  parts  being  zinc  and  lead.  The  total  extraction  is 
87%  of  the  contained  gold  and  silver. 

It  will  be  noticed  that,  whatever  the  grade  of  the  ore,  the  assay 
value  of  the  tailing  remains  uniform  (See  Fig.  94).  The  recovery 
not  only  applies  to  ore,  but  to  the  concentrate,  which  yields  98% 
when  treated  by  cyanidation.  In  the  practice  at  the  Guanajuato 
plant  the  low  extraction  is  due  to  decantation  of  the  slime  and 
could  be  increased  by  filter-pressing.  An  increase  also  would  result 
if  the  sand  were  more  finely  ground. 

The  itemized  cost  per  ton  of  ore  is  as  follows: 

Cyaniding :  Labor    $0.205 

Supplies,  including  cyanide,  etc. .  .  1.030 

Electric  power   0.130 

Assaying 0.060 

General  office  expense 0.070 

Plant  expense,  management,  etc ...  0.120 

1.615 

Crushing  and  sorting  ore 0.082 

Milling   0.550 


$2.247 


PART  V.     IRON 


PART  V.     IRON. 

68.     IRON  ORES. 

The  oxides  of  iron  occur  with  earthy  materials  and  never  in  a 
pure  state.  Only  those  are  called  iron  ores  that  contain  sufficient 
iron  to  make  the  recovery  of  the  metal  profitable.  Iron  ore  is  also 
used  as  a  flux  in  the  smelting  of  copper  and  lead  ores,  in  which  case 
the  iron  enters  the  slag  and  is  wasted.  The  three  kinds  of  iron  ore 
used  in  making  pig  iron  are :  the  hematites,  the  magnetites,  and  the 
carbonates. 

Hematite  (Fe203)  is  the  best  known  of  the  iron  ores.  It  contains, 
when  pure,  70%  iron.  When  the  proportion  of  contained  water  is 
low  it  is  called  red  or  brown  hematite,  while  the  hydrous  varieties 
are  called  soft  hematites  or  limonites,  although  the  name  limonite 
is  more  properly  applied  to  bog-iron  ore,  containing  20%  water. 
The  rich  varieties  (Lake  Superior  ores)  contain  58  to  64%  iron, 
while  the  limonites  run  as  low  as  50%.  By  roasting  the  ore  the 
water  can  be  mostly  driven  off  and  the  ore  improved  in  grade  and 
made  better  for  smelting.  Oolite  is  a  variety  that  exists  in  the  form 
of  grains  or  nodules  and  contains  silica  or  lime.  When  silicious,  as 
in  places  in  Alabama,  the  ore  is  well-nigh  worthless,  but  when 
limey,  as  in  the  Minette  district  of  Germany,  the  ore  is  self-fluxing, 
that  is,  the  lime  will  flux  the  silica  that  the  ore  contains.  It  then 
becomes  unnecessary  to  add  flux.  The  ore  of  these  districts  runs  as 
low  as  30  to  35%  iron,  yet  it  can  be  smelted  at  a  profit  because  of 
the  self-fluxing  property.  Red  hematite  iron  ore  is  the  most  desir- 
able. Most  of  the  Lake  Superior  deposits  are  of  this  variety,  and 
they  are  there  divided  into  two  classes :  the  hard  or  lumpy  ore  of  the 
Michigan  and  Wisconsin  ranges,  and  the  soft  ores  of  those  ranges, 
but  especially  of  the  Mesabi  range  in  Minnesota.  Among  the  ores 
of  the  United  States  we  should  also  mention  the  Alabama  beds,  and 
those  of  Colorado  and  Wyoming  in  the  West. 

Magnetite  (Fe304)  as  a  pure  mineral  contains  72.4%  iron.  It  is 
so  named  because  it  is  strongly  attracted  by  the  magnet,  whereas 
other  varieties  of  iron  ore  are  but  little  affected.  Magnetite  occurs 
in  large  deposits  in  Sweden,  and  in  various  parts  of  the  United 


274  THE    METALLURGY 

States.  While  some  of  the  beds  are  rich,  many  contain  no  more 
than  40%  iron  and  carry  so  much  silica  that  the  fluxing  and  smelting 
is  not  profitable.  Much  work  has  been  done  in  the  concentration 
of  these  ores,  both  in  Sweden  and  the  United  States.  In  New  Jersey 
extensive  beds  occur  that  have  been  utilized  by  Edison  for  the  pro- 
duction of  a  high-grade  ore  on  a  commercial  scale.  He  has  mined 
the  deposit,  crushed  and  concentrated  it,  and  made  it  into  briquettes 
that  contain  as  little  as  3.3%  silica  and  0.04%  phosphorus,  and  as 
high  as  67%  iron.  The  enterprise,  however,  could  not  continue  at  a 
profit  in  competition  with  foreign  ores  from  Cuba  and  Spain.  The 
extensive  plant,  so  ingeniously  devised  and  constructed  by  him,  is 
not  now  in  operation.  Some  of  the  New  York  beds  near  Lake 
Champlain  have  been  considered  valueless  on  account  of  the  pres- 
ence of  titanium,  it  having  been  asserted  that  this  element  produces 
an  infusible  sticky  slag.  This,  however,  has  been  proved  to  be  un- 
founded, and  it  should  not  prevent  their  use  as  a  source  of  iron. 
It  is  to  be  noted  that  the  famous  Iron  Mountain,  Missouri,  is  a 
deposit  containing  31%  iron  and  6%  titanium  oxide,  but  the  deposit 
is  not  now  worked.  In  Pennsylvania  the  Cornwall  beds  are  the 
most  important,  and  yield  a  pig-iron  carrying  not  more  than  0.04% 
phosphorus.  The  ore  runs  2.5%  sulphur,  and  about  half  of  this  is 
removed  by  kiln-roasting  before  smelting.  It  contains  also  copper, 
which  will  be  found  in  the  pig  to  the  extent  of  0.5  to  0.75%.  This 
does  not  matter  in  the  finished  product,  but,  if  the  pig-iron  is  made 
into  steel,  the  copper  causes  'hot-shortness'  or  brittleness  when  hot,* 
thus  causing  imperfections  when  rolled  into  shapes.  The  average  of 
ore  mined  is  40  to  42%  iron  and  20%  silica. 

Carbonate  ore,  siderite  (FeCO3),  as  a  pure  mineral  contains  48.3% 
iron.  The  varieties  are  spathic,  black  band,  clay-band,  or  clay  iron- 
stone. It  is  often  roasted  to  expel  the  moisture  and  carbon  dioxide 
before  going  to  the  blast-furnaces.  In  England  it  forms  the  well 
known  clay  iron-stone  of  the  Cleveland  district,  but  in  the  United 
States,  though  widely  distributed,  it  is  too  low  in  grade  to  be  used 
in  competition  with  the  abundant  rich  ores. 

69.     SMELTING  FOR  PIG-IRON. 

Outline  of  the  process. — The  operation  is  conducted  in  a  furnace, 
often  100  ft.  high,  filled  with  a  mixture  of  coke,  iron  ore,  and  lime- 
stone. Superheated  air  is  blown  in  at  the  bottom.  The  coke  is 
burned  to  maintain  a  high  temperature  in  the  furnace  and  to  reduce 
the  iron  in  the  ore  to  the  metallic  form  as  pig-iron.  The  pig-iron 
collects  at  the  hearth  or  bottom  of  the  furnace,  and  is  removed  from 


OF    THE    COMMON    METALS. 


275 


time  to  time.  The  gangue,  or  silicious  part  of  the  ore,  is  fluxed 
with  limestone,  and  produces  a  worthless  slag,  or  cinder,  which  is 
also  removed  (tapped)  as  it  accumulates  in  the  furnace. 

70.     BLAST-FURNACE  PLANT. 

Fig.  120  is  a  view  of  an  iron  blast-furnace  plant  for  the  manu- 
facture of  pig-iron  from  iron  ores.    In  the  foreground  is  a  cylindrical 


Fig.    120.      IRON   BLAST-FURNACE. 


276  THE    METALLURGY 

furnace-stack  100  ft.  high,  immediately  in  front  of  which  is  the 
forked  'down-comer'  (See  39,  Fig.  122),  a  large  pipe  that  conveys 
the  smoke  from  the  stack  near  the  top  downward  to  the  flue-system 


Fig.   121.     IRON  BLAST-FURNACE  WITH  AUTOMATIC  CHARGING. 

that  carries  it  away.  In  front  of  the  down-comer  we  see  the  inclined 
hoist  for  the  'stock'  or  the  materials  that  are  put  into  the  furnace. 
At  the  middle  of  the  illustration  are  the  four  cylindrical  'stoves',  as 
high  as  the  furnace,  used  for  pre-heating  the  air  blown  into  the  furn- 


OF    THE    COMMON    METALS.  277 

ace,  while  the  highest  stack  behind  them  draws  away  the  gas  from 
the  stoves.  In  front  of  the  four  stoves  is  the  blast-main,  a  pipe  5  ft. 
diam.,  by  which  the  air  is  conducted  to  the  furnace.  At  the  left  of 
the  stoves  is  the  building  (not  shown),  that  contains  the  vertical 
blowing  engines  by  which  air,  under  15-lb.  pressure,  is  delivered 
through  the  stoves  to  the  blast-furnaces.  Fig.  125  represents  a  verti- 
cal blowing  engine. 

The  general  arrangement  about  the  furnace  is  understood  from 
the  sectional  elevation  Fig.  121.  The  blast-furnace  100  ft.  high  is  at 
the  right.  It  is  served  by  an  inclined  hoist,  one  skip  of  which  is  in 
position  in  the  pit  ready  for  loading,  the  other  in  discharging  posi- 
tion at  the  top.  Within  the  stock-house,  at  the  left,  are  three  stan- 
dard-gauge tracks  on  three  levels.  The  upper  one  at  the  left  is  for 
the  hopper-cars  that  deliver  iron  ore  and  fluxes  to  sloping-bottom 
bins  beneath,  shown  in  section. 

The  second  one  leads  to  similar  bins  (not  shown  in  section),  and 
the  third  to  the  floor  on  the  ground-level.  On  this  level  is  a  charge- 
car,  electrically  driven,  with  a  weighing  attachment  that  can  be 
brought  to  any  bin  to  receive  a  weighed  amount  of  stock.  The  load 
is  then  transferred  to  and  discharged  into  the  skip.  In  case  of 
accident  to  the  charge-car,  or  any  trouble  at  bins,  the  furnace  can 
be  supplied  by  the  use  of  hand-barrows  or  buggies,  taking  the  stock 
from  the  piles  that  have  been  made  beneath  the  third  track.  Hoisting 
is  done  by  a  hoisting  engine  set  well  out  of  the  way  at  the  top  of 
the  stock-house.  Just  beyond  and  at  the  right  of  the  furnace  is  the 
cast-house  where  the  molten  iron  is  molded  into  pigs  when  a  cast 
is  made. 

71.     IRON  BLAST-FURNACE. 

Fig.  122  is  a  furnace,  shown  in  part  section,  and  part  elevation. 
It  is  circular  in  cross-section. 

Beginning  at  the  bottom  we  have  a  heavy  foundation  of  concrete 
and  fire-brick  upon  which  rests  the  hearth  15  and  columns  4  which 
support  the  upper  brick  work  that  constitutes  the  shaft  of  the 
furnace.  The  hearth  or  crucible  (14%  ft.  diam.  by  O1/^  ft.  deep), 
that  contains  the  molten  iron  and  slag,  extends  from  the  foundation 
to  a  height  slightly  above  the  tuyeres  22.  The  bottom  and  walls 
are  of  fire-brick,  and  near  the  bottom  is  the  iron-tap  27  through  which 
the  molten  pig-iron  is  withdrawn  when  a  quantity  has  accumulated. 
At  24  is  the  cinder-notch  or  tap  by  which  the  slag  or  cinder  is 
drawn  off.  The  crucible  is  surrounded  by  a  hearth-jacket  of  steel 
plates,  cooled  on  the  outside  by  sprays  of  water  that  play  against 


278 


THE    METALLURGY 


it,   cooling  and  protecting  it  and  the  brick-work  lining  from  the 
corrosive  action  of  the  molten  slag  inside.     Air^  under  a  pressure 


Fig.    122.      IRON  BLAST-FURNACE,    DETAILED   SECTION. 

of  5  to  14  lb.  to  the  square  inch  enters  through  the  tuyeres  21, 
which  have  projecting  nozzles  22.  Care  is  taken  to  withdraw  the 
slag  before  it  reaches  the  level  of  the  tuyeres,  for  it  would  enter  the 


OF    THE    COMMON    METALS.  279 

openings  and  close  them.  Of  these  tuyeres  there  are  six.  The  air 
is  supplied  through  the  tuyere-stocks  33  from  the  bustle-pipe  13, 
which  encircles  the  furnace,  and  connects  with  the  blast-main 
supplying  air  at  the  temperature  of  a  red  heat  from  the  stoves. 
The  bustle-pipe,  tuyere,  and  tuyere-stock  are  shown  in  the  section 
Fig.  121.  The  bosh,  or  that  part  of  the  furnace  that  widens  from 
14  ft.  6  in.  at  the  hearth  to  22  ft.  in  a  vertical  distance  of  13  ft. 
is  also  shown.  It  is  in  the  region  of  the  bosh  that  the  formation  of 
the  slag  occurs,  and  the  brick-work  of  the  bosh  is  subject  to  a 
slagging  and  scouring  action  that  tends  to  attack  and  destroy  it. 
To  prevent  this,  hollow  water-cooled  bosh-plates  14  are  laid  in  the 
brick-work  of  the  bosh,  making  rings  around  the  furnace  at  nearly 
every  two  feet  vertically.  The  slag  cuts  into  the  brick-work  nearly 
as  far  as  the  inner  ends  of  these  plates,  but  the  circulation  of  water 
within  them  protects  the  adjacent  brick-work  from  deeper  corrosive 
action. 

The  shaft,  or  main  brick-work  structure  of  the  furnace,  is  carried 
by  the  cast-iron  mantel  5  resting  upon  the  columns  4.  It  extends 
from  the  top  of  the  bosh  to  the  throat  at  9.  The  upper  part  of  the 
furnace  is  closed  by  a  bell  47  (Fig.  121  shows  a  double  bell),  and 
the  gas  escapes  at  the  side  through  the  down-comer  39.  The  in-wall 
69  is  of  fire-brick,  while  the  main  portion  is  of  common  brick  and 
is  sheathed  with  a  shell  46  of  steel  plates.  When  in  operation 
the  furnace  is  kept  full  to  a  level  just  below  the  outlet  to  the 
down-comer.  This  level  is  known  as  the  stock-line,  and  the  furnace 
at  this  point  is  15  ft.  diam.  As  the  stock  smelts  and  sinks,  charges 
are  introduced  and  the  stock-line  is  maintained  at  this  level. 

Ordinarily  the  top  of  the  furnace  is  kept  closed  by  the  conical 
bell  47  which  is  suspended  from  the  ends  of  the  counterweighted 
beams  55.  The  bell  closes  the  bottom  of  a  circular  hopper  48,  into 
which  the  charge  in  this  particular  furnace  is  supplied  by  buggies 
brought  up  by  elevator  to  the  upper  or  charge-floor  of  the  furnace 
or  'tunnel-head,'  as  it  is  called.  To  drop  a  charge  into  the  furnace, 
the  outer  end  of  the  lever  is  raised  by  the  piston-rod  and  piston 
of  the  air-cylinder  60.  The  bell  thus  lowered  permits  the  charge 
to  slide  into  the  furnace,  after  which  it  is  immediately  raised  to 
close  the  opening  and  stop  the  outward  rush  of  smoke  and  gas  that 
mainly  escape  through  the  hood  61.  The  gas,  containing  dust  from 
the  charge,  passes  off  by  the  down-comer  39  to  the  dust-catcher  40 
(where  a  part  of  the  dust  settles),  and  by  the  goose-neck  pipe  41 
to  an  underground  flue  that  leads  to  the  stoves  and  boilers  where 
the  gas  is  burned.  Rising  from  the  down-comer  is  the  bleeder  37? 


280  THE    METALLURGY 

that  is  used  when  it  is  desired  to  relieve  the  top  pressure  of  the 
gas  rising  from  the  charge.  It  is  occasionally  used.  At  many 
furnaces  the  stock  is  raised  in  hand-barrows  or  charge-buggies  to 
the  furnace-top  or  tunnel-head  51  by  means  of  a  platform  hoist. 
In  Fig.  120  is  seen  at  the  extreme  right,  the  frame  of  the  hoist. 

In  Fig.  121  is  shown  the  more  recent  method  of  charging  with  the 
inclined  hoist.  A  double  bell  is  used  to  prevent  the  escape  of  the 
gas.  The  charge  is  dropped  from  the  hoist  into  the  upper  hopper, 
where  it  is  retained  until  the  lower  hopper  is  empty.  The  smaller 
upper  bell  is  then  lowered  and  the  charge  slides  from  the  upper 
into  the  lower  hopper,  while  the  upper  bell  is  closed.  The  hopper  is 
then  ready  to  take  another  charge.  The  charge  in  the  lower  hopper, 
when  needed  is  dropped  into  the  furnace  by  lowering  the  lower 
bell.  It  slides  outwardly  to  the  walls,  forming  a  ring  or  ridge, 
the  stock  in  the  middle  being  a  little  lower  than  at  the  sides. 
Some  coarse  lumps  roll  toward  the  middle,  so  that  part  of  the 
charge  is  more  open  than  at  the  walls.  Just  beyond  and  below 
the  lower  bell  is  noticed  the  oval  outlet  to  the  down-comer.  The 
stock-line  must  be  kept  below  this. 

The  skips  of  the  hoist  run  in  balance  and  are  charged  as 
follows:  The  charge-car  on  the  ground  level  is  run  to  the  chute 
of  an  iron-ore  bin  to  receive  the  required  weight  of  ore.  It  is 
moved  to  the  limestone-bin  beyond  to  get  the  needed  quantity  of 
limestone,  and  then  to  the  skip  standing  below  in  the  charge-pit 
where  it  is  discharged.  The  skip  is  next  hoisted  and  dumped, 
while  the  empty  one  is  in  position  to  take  the  load  of  coke.  After 
elevating  the  fuel,  a  charge  of  ore  and  flux  next  goes.  These 
charges  alternate  in  the  furnace  and  form  layer  upon  layer. 

The  dimensions  of  a  blast-furnace  are  limited.  The  considerations 
are  as  follows:  The  hearth  should  be  not  more  than  15  ft.  diam. 
lest  the  blast  fail  to  properly  penetrate  to  the  center  and  maintain 
intense  combustion  there.  The  slope  or  angle  of  the  bosh-wall  must 
be  such  as  to  give  proper  support  to  the  charge,  which  rests  upon 
it,  and  yet  allow  the  solid  coke  to  slip  down.  The  height  is 
limited  to  the  height  of  the  smelting  zone.  These  conditions  limit 
the  diameter  of  the  bosh  to  22  ft.  From  the  top  of  the  bosh  the 
stack-wall  must  decrease  in  diameter  to  the  throat  to  give  room  for 
the  descending  charge  to  swell  by  reactions  that  occur  in  the  down- 
ward progress.  This  leaves,  at  the  throat,  a  diameter  suitable  for 
the  proper  distribution  of  charge.  Furnaces  have  been  built  higher 
than  100  ft.,  but  such  height  has  been  found  to  be  excessive, 


OF    THE    COMMON    METALS. 


281 


especially  for  fine  ores;  and  the  best  practice  calls  for  90  ft.  or 
less. 

72.     THE  STOVES. 

The  efficient  operation  of  an  iron  blast-furnace  requires  that 
the  air  entering  at  the  tuyeres  be  heated,  generally  to  the 
temperature  of  a  red  heat  (500  to  750°C).  To  do  this  the  furnace 
is  equipped  with  three  or  four  (as  in, Fig.  120)  regenerative  fire- 


Fig.   123.      COWPER  HOT-BLAST   STOVE    (ELEVATION). 


brick  stoves  80  ft.  high  and  14  ft.  diam.  The  Cowper  stove  (of  Fig. 
123  and  124),  for  example,  consists  of  a  tight  shell,  like  a  boiler 
shell  of  steel  plates,  lined  with  fire-brick,  and  containing  a  checker- 
work  of  bricks  of  special  shape  laid  in  open  order  so  as  to  have 
numerous  openings  or  passages  from  the  top  to  the  bottom  of  the 
stove.  The  gas  from  the  furnace,  containing  24%  CO,  which  in 


282 


THE    METALLURGY 


burning  supplies  the  heat,  flowing  along  the  underground  flue  from 
the  goose-neck  before  mentioned,  enters  the  stove  at  </,  while  air  is 
admitted  at  a,  the  two  gases  mingling  and  burning  in  the  vertical 
circular  flue  f,  and  heating  the  checker-work  in  their  passage  to 
the  valve  s  as  shown  by  the  arrows.  The  gas  thence  passes  by  an 
underground  flue  to  a  tall  stack,  200  ft.  high,  shown  behind  the 
stoves  in  Fig.  120.  Gas  having  burned  in  this  way  a  half  hour, 
the  stove  becomes  heated.  The  valves  g,  a,  and  s  are  closed,  and  the 
cold-air  valve  at  c  near  the  bottom  of  the  stove  is  opened,  admitting 
air  under  pressure  from  the  blowing-engine,  the  hot-air  valve  //, 


Fig.   124.      COWPER  HOT-BLAST   STOVE    (PLAN). 

at  the  bottom  of  the  flue  /,  being  at  the  same  time  opened.  The 
air  rises  through  the  hot  checker-work,  descends  the  flue  /  and 
passes  out  at  d  to  the  brick-lined  hot-blast  main,  and  to  the  furnace. 

Meanwhile  the  gas  has  been  turned  into  the  other  stoves,  and 
is  heating  them  in  the  same  way.  After  the  blast  has  been  received 
in  the  first  stove  a  half  hour  it  is  turned  into  the  next  heated 
stove,  and  so  on.  The  extensive  surface  of  the  checker-work  serves 
to  absorb  a  large  amount  of  heat  from  the  burning  gas,  and  to  impart 
the  heat  subsequently  to  the  blast-air.  It  will  be  noticed  that  the 
air  enters  at  the  coldest  and  leaves  at  the  hottest  part  of  the 
stove.  It  flows  in  a  direction  opposite  to  that  which  the  burning 
gas  does,  thus  insuring  the  maximum  rise  in  temperature.  In 
the  course  of  half,  an  hour,  the  hot  air,  leaving  the  stove,  falls 
at  least  100 °C.  in  temperature. 

Not  all  the  gas  from  the  furnace  is  needed  for  heating  the 
stoves,  and  a  portion  is  burned  under  the  boilers  of  the  plant 
for  making  steam  for  power.  The  amount  of  steam  thus  available 
is  sufficient  to  run  the  blowing-engines,  hoisting-mechanism,  and 


OF   THE    COMMON    METALS.  283 

all  machinery  belonging  to  the  furnace.  At  some  plants  the  surplus 
gas,  after  the  stoves  have  been  supplied,  has  been  cleaned  from 
dust  and  used  in  gas-engines.  Power  can  be  gained  in  this  way 
and  is  more  available  for  the  rolling  mill  or  other  purposes. 

Formerly  the  air  was  heated  in  iron-pipe  stoves,  the  air 
circulating  through  a  nest  of  pipes  inclosed  in  a  furnace  or  brick 
heating-chamber.  These  are  no  longer  used,  but  have  been 
supplanted  by  the  regenerative  stove  already  described. 

73.     BLOWING  ENGINES. 

Fig.  125  gives  a  view  of  a  vertical  blowing-engine,  having  an 
air-cylinder  at  the  top  7  ft.  in  diameter  and  of  a  5-ft.  stroke.  The 
cylinder  displaces  385  cu.  ft.  air  per  revolution,  or  15,400  cu.  ft. 
per  minute,  at  15-lb.  pressure.  The  air-admission  and  discharge 
valves  are  arranged  to  operate  positively,  and  to  open  and  close 
exactly  at  the  right  moment.  Enough  of  these  engines  are  installed 
to  supply  the  amount  of  air  required  by  the  furnace. 

74.     OPERATION  OF  THE  FURNACE. 

The  ore  is  dumped  into  the  furnace  with  45  to  60%  of  the  weight 
of  coke,  and  the  limestone  needed  to  form  the  predetermined  slag. 
The  furnace  should  be  at  least  65  ft.  high,  and  is  now  built  80  to 
100  ft.  high.  It  is  kept  full  of  stock,  and  the  combustion  of  the 
coke  is  supported  by  the  air  introduced  at  high  pressure  at  the 
tuyeres.  As  smelting  progresses,  the  coke  burns,  the  slag  and 
iron  produced  from  the  charge  is  withdrawn,  and  the  surface  or 
stock-line  sinks.  Thus  the  removal  of  molten  products  below  and 
the  addition  of  fresh  stock  above,  causes  the  greatest  production 
of  heat  next  the  tuyeres,  where  the  coke  largely  burns,  the 
temperature  decreasing  toward  the  stock-line.  The  actual  melting 
zone,  or  zone  of  fusion,  extends  upward  through  the  bosh-region. 
The  most  intense  combustion  occurs  within  4  ft.  of  the  tuyeres, 
where  an  excess  of  air,  driven  in  under  high  pressure,  burns  the 
coke  to  carbon  dioxide.  In  the  reaction,  the  CO2  may  be  said  to 
dissolve  the  carbon,  it  being  as  follows: 

CO2  +  C  =  2CO 
97,000     2  x29,000  =  -  39,000 

The  reaction  being  endothermic,  lessens  the  temperature  in  this 
second  region.  The  temperature  is  so  high,  however,  that  the 
glowing  coke  reduces  to  iron  any  iron  oxide  that  descends  as  far 
down  in  the  furnace  as  this  region.  The  rising  gas  consists  of  CO 


284  THE    METALLURGY 

and  N.  The  carbon  monoxide,  in  the  ascent,  reduces  the  iron  oxide 
to  iron.  On  reaching  the  lower  part  of  the  furnace  the  iron,  with 
carbon  taken  from  the  CO,  forms  pig-iron.  The  iron  takes  up 
silica,  phosphorus,  and  sulphur  from  the  earthy  constituents  of  the 
charge.  The  carbon  amounts  to  3  or  4%,  and  the  other  impurities 
to  2%,  the  weight  of  the  product.  It  is  the  5%  of  the  metalloids 
present  in  the  pig  that  makes  it  fusible. 

We  have,  therefore,  in  the  blast-furnace  beginning  from  above, 
three  zones: 

(1)  The  zone  of  preparation,   where  C02  is  driven  from  the 
limestone  and  moisture  from  the  charge. 

(2)  The  zone  of  reduction,   where  the  CO  of  the  rising  gas 
reduces  the  iron  ore,  first  to  the  ferrous  form,  then  to  iron,  and 
where  the  iron  in  a  spongy  or  open  form  absorbs  carbon  from  the 
reducing  gas. 

(3)  The  zone  of  fusion,  where  the  temperature  of  the  furnace 
is  high  and  the  slag  is  formed,  the  iron  at  the  same  time  absorbing 
silicon,  phosphorus,  and  sulphur. 

In  the  upper  zone  of  the  furnace,  the  carbon  dioxide  of  the 
limestone  is  expelled,  leaving  quicklime  (CaO)  ready  for  fluxing 
the  gangue  or  waste  matter  of  the  charge.  The  reaction  is 
endothermic,  lessening  the  heat  of  the  escaping  gases.  The 
quantity  of  limestone  needed  to  form  a  suitable  slag  is  calculated 
in  advance. 

The  gas  varies  in  composition,  but  commonly  contains  61% 
nitrogen,  from  10  to  17%  CO2,  and  22  to  27%  CO.  The  carbon 
monoxide  is  the  combustible  constituent  of  gas  that  produces  the 
heat  when  the  gas  is  burned  in  the  stoves  and  boilers. 

In  the  foregoing  description  it  has  been  assumed  that  coke  is 
the  fuel  employed  as  is  true  in  most  cases.  Anthracite  coal  has 
been  used  when  cheap  enough  to  compete  with  coke,  but  even  then 
a  more  satisfactory  result  is  obtained  when  coke  forms  part  of  the 
charge.  Furnaces  in  which  a  part  of  the  fuel  is  anthracite,  are 
called  anthracite  furnaces,  but  the  name  is  somewhat  misleading. 
Other  furnaces  use  charcoal  exclusively.  Charcoal  is  supposed  to 
give  an  iron  of  great  toughness  that  is  particularly  valuable  for 
cast-iron  car-wheels,  and  other  castings  requiring  toughness.  Its 
superiority  over  other  kinds  having  the  same  constitution  has  been 
widely  disputed,  but  there  is  testimony  in  favor  of  charcoal-iron. 

Thirty  years  ago,  blast-furnace  practice  was  regulated  by  rule- 
of -thumb  methods.  They  were  "born  of  a  bigoted  belief,  on  the 
part  of  ignorant  furnace-men,  that  particular  ores  and  fuel  could 


OF    THE    COMMON    METALS. 


285 


be  worked  in  a  furnace  only  on  special  lines,  and  that  it  was 
impious  to  drive  a  furnace  faster  than  a  certain  rate  established  by 
time-worn  tradition."  In  1879,  certain  experiments  made  at  the 
Edgar  Thompson  Steel  Works,  Pittsburg,  Pa.,  showed  that  it  was 
possible  to  increase  the  output  of  a  furnace  enormously  by  increasing 
the  air-supply.  It  was  also  found  that  the  amount  of  air,  not 
the  pressure,  determined  the  rapidity.  Under  the  new  system  it 


Fig.   125.     BLOWING  ENGINE. 

was  thought  necessary  to  make  a  steep-angle  bosh  (80  degrees) 
resembling  that  in  Fig.  122  more  than  that  in  Fig.  121.  With  the 
more  rapid  driving,  reduction  decreased,  and  the  slag  contained  more 
iron.  To  secure  the  reduction,  the  fuel  had  to  be  kept  high, 
using  one  ton  of  coke  per  ton  of  pig-iron  produced,  and  where 
coke  was  expensive  this  was  a  serious  matter.  E.  C.  Potter  at  the 
Illinois  Steel  Works,  South  Chicago,  showed  that  by  reducing  the 


286  THE    METALLURGY 

bosh-angle  to  75%  and  using  somewhat  less  blast,  it  was  possible 
to  cut  the  coke  consumption  from  2240  Ib.  to  1800,  or  even  1750, 
per  ton  of  pig  produced.  Furnaces  with  large  hearths  were  then 
built,  which  also  increased  capacity. 

Blowing-in. — The  furnace  is  first  dried  several  days  by  a  wood 
fire  in  the  crucible.  The  lower  part,  halfway  up  the  bosh,  is 
filled  with  cord-wood.  Upon  this  is  placed  a  heavy  bed  of  coke 
with  limestone  to  flux  the  coke-ash,  followed  by  successive  layers 
of  the  normal  charge  of  coke,  with  gradually  increasing  amounts 
of  ore  and  limestone,  and  decreasing  quantities  of  slag,  until  the 
normal  charge  of  ore  and  flux  is  reached.  The  wood  is  ignited  at 
the  tuyeres,  and  a  weak  blast  of  air  supplied.  The  pressure  is 
gradually  increased  during  24  hours,  and  the  furnace,  becoming 
entirely  filled  during  this  time,  the  regular  pressure  is  reached. 

Regular  operation. — -The  old  way  of  charging  the  furnace,  by 
hand,  is  as  follows:  Ores,  flux,  and  fuel  (the  'stock')  is  brought 
in  buggies  from  the  stock-house  to  the  scales,  weighed,  hoisted  to 
the  top  of  the  furnace,  or  tunnel-head,  wheeled  by  top-fillers  to 
the  bell,  and  dumped  evenly,  the  fuel  separately,  the  ore  and  'stone' 
(limestone)  together.  The  furnace  is  kept  full  to  a  level  just  below 
the  outlet  of  the  down-comer.  In  automatic  charging,  as  indicated 
in  Fig.  121,  the  stock  is  charged  to  the  skips,  as  has  been  described, 
and  is  hoisted  to  the  furnace-top  and  dumped  into  the  double  hopper. 
No  top-fillers  are  needed,  and  the  bells  are  often  operated  from 
the  ground-level,  so  that  no  attendant  is  needed  at  the  tunnel-head. 

75.     DISPOSAL  OF  SLAG  OR  CINDER. 

On  account  of  the  low  specific  gravity,  slag  floats  upon  the 
iron.  The  iron  occupies  the  lower  part  of  the  crucible,  and 
accumulates,  until  it  reaches  the  tuyeres  when  it  should  be  drawn 
off.  Every  1%  to  2  hours,  the  plugged  cinder-notch  is  pierced  with 
a  pointed  steel  rod,  and  the  cinder  above  the  level  allowed  to  flow 
out.  It  flows  along  a  cast-iron  launder  a  distance  of  15  to  30  ft. 
and  falls  into  a  14-ton  slag-car  standing  on  a  track  below.  When 
loaded  the  car  is  hauled  by  a  locomotive  to  the  dump  that  may 
be  a  mile  away,  and  the  contents  of  the  ladle  is  poured  out  at  the 
side  of  the  track.  The  track  is  gradually  raised  and  moved  outward 
toward  the  edge  of  the  dump  as  it  grows. 

A  large  coke  furnace,  yielding  500  tons  of  pig  daily,  when 
smelting  ores  of  good  grade,  produces  daily  300  tons  of  cinder. 
In  smelting  silicious  ores,  the  quantity  of  slag  may  be  twice  as 
great. 


OF   THE    COMMON    METALS. 


287 


76.     DISPOSAL  OF  PIG-IRON. 

Every  four  to  six  hours  the  metal  is  tapped  from  the  furnace, 
50  tons  at  a  time.  The  flow  is  started  by  a  pointed  steel  bar 
which  is  driven  through  the  clay-plugged  iron-notch  or  tap- 
hole.  A  clay-lined  launder  conducts  the  flow  to  ladles  similar 
to  the  cinder  car.  Distant  from  the  furnace  10  to  15  ft.  is 
a  cross-channel  made  in  the  sand.  The  slag  that  floats  on  the 
stream  of  iron  is  diverted  into  the  cross-channel  with  a  skimmer. 
An  iron  plate  is  placed  across  the  flowing  stream  in  such  a  way  as 
to  permit  the  heavy  iron  to  flow  beneath,  while  the  light  slag  is 
diverted.  A  charcoal  furnace,  having  an  output  of  100  tons  per 
day,  produces  so  little  cinder  that  none  is  tapped  at  the  cinder- 
notch,  but  flowing  out  with  the  metal  it  is  skimmed  as  above 
described.  It  is  run  outside  the  cast-house  upon  the  ground,  is 
allowed  to  cool,  and  is  then  broken  up,  and  carted  away.  The 


Fig.   126.      HEYL  AND  PATTERSON  PIG-CASTING  MACHINE. 


iron  contained  in  the  ladles  is  called  'direct  metal'  and  is  taken 
to  the  steel-works  and  used  in  molten  form. 

In  practice  elsewhere,  the  iron  is  cast  in  molds  in  the  sand 
of  the  floor  of  the  cast-house.  The  floor,  40  ft.  wide  and  80  to 
150  ft.  long,  consists  of  the  sand  in  which  the  depressions  are 
molded  and  connected  by  a  main-channel  or  runner,  which  receives 
the  molten  iron  flowing  from  the  furnace.  The  molds  fill  successively 
and  form  pigs  of  iron  weighing  150  Ib.  each. 

To  reduce  the  cost  of  handling  the  pig  metal,  and  to  give  a 
product  smooth  and  free  from  adhering  sand,  casting  machines  have 
been  introduced  in  modern  plants.  Fig.  126  is  a  Heyl  &  Patterson 
pig-casting  machine,  consisting  of  an  endless-chain  conveyor 
composed  of  a  series  of  molds,  each  capable  of  holding  120  Ib.  iron. 
The  iron,  brought  from  the  furnace  in  a  large  ladle  shown  at  d, 
is  poured  into  the  molds  as  they  travel  slowly  along.  The  pig- 
iron  chills  quickly,  and  by  the  time  it  reaches  the  discharge  end,  it 
consists  of  solid  pigs  of  iron  and  drops  into  the  railroad  car  that 


288  THE    METALLURGY 

is  placed  in  position  to  receive  it.  The  molds  on  their  return, 
inverted,  take  at  c  a  spray  of  whitewash,  the  water  of  which  quickly 
dries  by  the  heat  of  the  mold.  It  leaves  a  coating  of  lime  inside 
that  prevents  the  iron  from  adhering.  Mechanical  casting  has  the 
advantage  over  casting  in  sand-molds  that  it  does  away  with  the 
hot  and  severe  work  of  breaking  and  handling  the  pigs.  In  hot 
weather  the  work  can  hardly  be  borne,  and  there  always  is  difficulty 
in  getting  or  keeping  the  men. 

Blowing-down. — When  a  furnace  is  to  be  put  out  of  blast, 
charging  is  stopped,  and  a  layer  of  coke  is  added  for  the  last 
charge.  With  continued  blast  the  stock-line  descends  and  the 
operation  progresses  as  long  as  iron  and  cinder  can  be  tapped  out, 
the  blast  being  gradually  diminished.  Finally  the  blast  is  stopped, 
and  the  remainder  of  the  content  is  withdrawn  through  a  hole 
broken  in  the  brick-work  near  the  bottom. 

77.     CHEMICAL  REACTIONS  OF  THE  IRON  BLAST-FURNACE. 

A  blast-furnace  may  be  likened  to  an  immense  gas-producer 
in  which  there  is  a  column,  70  ft.  high,  of  alternate  layers  of 
coke,  iron  ore,  and  flux.  The  column  ranges  in  temperature  from 
a  heat  that  shows  no  color  at  the  throat,  to  a  \vhite  heat  at  the 
tuyeres. 

The  hot  air  of  the  blast,  entering  at  the  tuyeres,  strikes  the 
white-hot  coke  with  the  immediate  formation  of  CO2  followed  by 
an  instananeous  reduction  to  CO.  The  air  therefore  need  only  burn 
the  fuel  to  CO  as  indicated  by  the  following  reaction : 

(1)     C  -f  0  =  CO  =  29,000  Calories 

Per  pound  of  carbon  burned  2415  pound-calories  are  generated. 
Since  23%  of  air  is  oxygen,  and  at  the  sea-level  one  pound 
of  air  equals  12.38  cu  ft.,  we  have  ±|  X  y~  =  72  cu.  ft.  of  air 
per  pound  of  carbon,  or  61  cu.  ft.  per  pound  of  coke  of  85% 
carbon. 

Fig.  127  shows  graphically  the  chemical  reactions  under  a  set 
of  conditions  assumed,  while  the  temperature  and  places  where  the 
reactions  take  place  are  shown  in  the  section  of  the  furnace  at 
the  extreme  left  in  the  diagram.  To  produce  a  ton  of  pig-iron 
(2240  Ib.)  there  is  to  be  used  3520  Ib.  of  60%  iron  ore  containing 
3020  Ib.  Fe2O8,  1888  Ib.  coke,  and  1010  Ib.  limestone. 

At  the  tunnel-head,  the  iron  ore  (Fe203)  plunges  into  an 
atmosphere  of  24%  CO,  16%  CO2,  and  60%  of  N  at  a  temperature 
of  260°C.  Reduction  of  ferric  oxide  to  Fe3O4  by  the  CO  begins 
thus : . 


OF    THE    COMMON    METALS. 


289 


(2)     3Fe203  +  CO  =  2Fe3O4 


CO, 


3x199,400    29,000    2x268,500    97,000  =  + 11,400   Cal. 

The  reaction  is  completed  at  a  temperature  of  450°  C.  when  the  ore 
has  reached  a  depth  of  10  ft.,  shown  at  2  of  the  diagram.  During 
this  period,  the  peculiar  reaction  resulting  in  carbon-deposition, 

(1)      (2)      (3)      (4)     (5)      (6)    (7)(8)(9)      (10)      (11)          (12)          (13) 


Fig.   127.      CHEMICAL  REACTIONS  OF  BLAST-FURNACE. 


begins,  caused  by  the  reaction  of  the  gas  on  the  ore,  forming  a 
deposit  of  soot  or  carbon  in  the  pores. 

(3)  2Fe203  +  SCO  =  7C02  -^e  +  C. 

Continuing  the  descent  the  ore  undergoes  further  reduction. 
At  a  depth  of  19  ft.  and  a  temperature  of  850°C.,  the  Fe304  formed, 
as  shown  above,  has  become  further  reduced  to  FeO,  as  indicated 
in  column  4,  the  reaction  being  as  follows : 

(4)  Fe3O4  +  CO  =  3FeO  +  CO2 

265,800  29,000  3x66,400  97,000—41400  Cal. 
The  FeO  thus  formed,  impregnated  with  carbon  (See  column  7) 
descends  with  little  change,  until  at  a  depth  of  26  ft.  and  at  a 
temperature  of  700° C.,  the  CO  of  the  gas  reacts  upon  it,  and  spongy 
iron  begins  to  form.  The  reaction  is  complete  at  800°C.,  and  at 
a  depth  of  32  ft.,  and  is  as  follows : 

.(5)     FeO  +  C02  =  Fe  +  C02       /6d0 

66,400     29,000     97,000  =  + frtfte  Cal. 


290  THE    METALLURGY 

In  the  passage  downward  the  limestone  gradually  loses  its  C02, 
and  at  this  point  the  expulsion  is  complete,  as  indicated  at  column 
8.  The  quicklime,  column  9,  thus  formed,  unites  at  the  zone  of 
fusion  and  fluxes  the  silica  of  the  charge.  From  the  depth  of 
19  ft.  to  the  depth  at  which  all  carbon  dioxide  is  expelled,  that  is 
between  the  temperatures  550°  and  880°C.,  the  C02  reacts  upon 
coke,  dissolving  it  according  to  the  following  reaction : 

(6)  CO2  +  C  =  2CO 

97,000     2x29,000  =     -  39,000  Cal. 

Thus  heat  is  absorbed  from  the  gas,  and  coke  is  consumed.  The 
coke,  however,  as  can  be  seen  from  column  6,  remains  but  little 
changed  until  it  reaches  the  region  of  the  tuyeres. 

Below  the  32-ft.  level  at  800°C.,  reactions  practically  cease,  the 
chief  action  now  being  a  reduction  of  a  small  amount  of  FeO, 
left  undecomposed  by  the  CO.  This  is  gradually  reduced  (See 
column  4)  by  the  glowing  coke  as  follows : 

(7)  FeO-{-C  =  Fe  +  CO 

66,400  29,000  =    -  37,400 

Silicon  having  less  affinity  for  oxygen  than  carbon  at  a  high 
temperature  is  formed  from  the  reduction  of  the  silica,  and  as  a 
metalloid  enters  the  pig-iron  after  the  following  reaction : 

(8)  Si02  +  2C  =  2CO  +  Si. 

Below  the  32-ft.  level,  the  temperature  rises  gradually  and  uniformly 
until  the  intense  combustion  at  the  tuyeres  produces  1500°C.  as  a 
maximum. 

Of  the  air  entering  the  furnace,  77%  is  nitrogen,  and  of  the 
escaping  gas  60%,  thus  showing  nitrogen  to  be  by  far  the  largest 
constituent  present.  As  is  shown  in  column  12,  nearly  three  tons 
of  nitrogen  pass  through  the  furnace  for  each  ton  of  pig-iron 
produced.  At  the  high  temperature  of  the  lower  part  of  the  furnace, 
potassium  cyanide  is  formed,  the  potash  of  the  coke-ash  uniting 
with  carbon  and  nitrogen  to  form  the  salt.  It  decomposes  before 
the  top  of  the  charge  is  reached. 

Referring  to  columns  10  and  11,  we  note  that  the  C02  formed 
so  freely  at  the  tuyeres,  is  at  once  (column  10)  changed  to  CO. 
The  carbon  monoxide  rises  unchanged  until  it  reaches  the  32-ft. 
level,  when  it  begins  to  act  on  the  iron  oxides  with  the  formation 
of  C02.  The  carbon  dioxide  from  this  source  united  with  that 
from  the  limestone  is  the  total  of  the  escaping  CO2  gas. 

Sulphur  occuring  as  FeS  in  the  coke  and  as  pyrite  in  the  ore, 
is  speedily  driven  off  by  the  heat  of  the  furnace,  giving  FeS.  The 


OF    THE    COMMON    METALS.  291 

sulphur  of  the  FeS  is  taken  up  by  the  quicklime,  and  enters  the 
slag  as  calcium  sulphide  according  to  the  following  reaction : 

(9)     FeS  +  CaO  +  C  =  CaS  +  Fe  +  CO. 

Thus  it  is  separated  from  the  iron  upon  which  it  would  have  an 
injurious  effect. 

78.  CALCULATION  OF  AN  IRON-FURNACE  CHARGE. 

To  be  able  to  calculate  a  blast-furnace  charge  is  of  prime 
importance  to  the  young  metallurgist.  His  superior  has  often  not 
the  time  or  disposition  to  impart  the  knowledge  to  him.  The 
accompanying  charge-sheet  illustrates  the  method  of  estimating 
a  charge  for  a  furnace  using  coke  with  an  assumed  composition  of 
pig-iron  and  slag.  We  wish  to  smelt  Mesabi  ore,  which  alone  gives 
too  compact  a  charge  for  smelting.  We  accordingly  add  to  it  lumpy 
ore.  Since  we  are  using  coke  that  contains  sulphur  and  since  we 
can  remove  the  sulphur  from  the  pig-iron  by  producing  a  limey 
slag,  we  prescribe  the  addition  of  lumpy,  silicious,  Maxwell  ore,  with 
the  silica  of  which  we  make  a  slag  of  the  composition  given  on 
the  charge-sheet,  containing  a  high  content  of  lime  to  insure  the 
removal  of  the  sulphur,  as  indicated  in  equation  (9)  above. 

While  the  iron  ores  used  in  the  large  iron  districts  contain  little 
sulphur,  the  coke  may  have  0.25  to  2%  or  an  average  of  1%.  The 
sulphur  must  not  enter  the  iron,  and  by  using  a  slag  containing 
plenty  of  lime,  and  hot  furnace,  the  sulphur  may  be  forced  into 
the  slag.  Thus  for  a  Lake  Superior  iron  ore,  using  coke,  a  suitable 
slag  had  the  composition  SiO2  35.1%,  A1203  14.2%,  CaO  22.5%, 
and  MgO  22.4%.  The  iron  contained  Si  1.4%  and  S  0.05%.  In 
another  instance,  in  a  fairly  hot  furnace,  a  slag  of  Si02,  36.1%,  A12O3 
12.9%,  CaO  41.7%,  and  MgO  7.3%,  took  up,  principally  from  the 
coke,  1.6%  S,  and  produced  an  iron  containing  Si  2.15%  with  only 
0.02%  sulphur.  The  slag  was  quite  clean,  containing  0.54%  Fe. 

We  will  now  enter  the  names  of  the  two  ores  and  the  weights 
(1000  Ib.  each)  as  well  as  the  percentage  composition,  in  the  two 
first  lines  of  the  charge-sheet,  and  fill  out  the  corresponding 
weights  of  the  silica  and  other  constituents.  On  the  next  lines, 
in  the  same  way,  we  enter  the  limestone  and  coke,  and  tabulate  the 
percentage  composition  of  the  limestone. 

In  the  column  marked  'Fe'  we  find  the  total  of  iron  to  be 
Ib.,  approximately  equivalent  to  1200  Ib.  of  the  92.3%  pig-iron 
given  on  the  charge-sheet.  From  the  experience  in  smelting  such 
ore  we  assume  that  one  ton  of  coke  is  sufficient  for  one  ton  of  pig- 


292 


THE    METALLURGY 


iron  produced,  so  that  for  the  1200  Ib.  of  pig  we  need  1200  Ib.  of 
coke,  and  we  insert  this  in  the  table. 

The  coke  carries  11%  ash  that  by  analysis  is  known  to  contain 
40%  SiO2,  9%  Fe,  9%  CaO,  and  32%  A12O3.  The  corresponding 
Si02  in  the  coke  will  be  11%  of  the  40%  of  the  ash,  or  4.4%,  and 
the  percentages  in  Fe,  CaO,  and  A1203  will  be  respectively  1.0% 
(accurately  0.99%),  1.0%,  and  3.5%.  The  percentage  of  sulphur  as 
reported  by  the  chemist  is  based  upon  the  weight  of  the  coke  and 
is  written  directly  with  the  other  percentages.  The  problem  is  to 
determine  how  much  limestone  is  to  be  added  to  obtain  a  slag  of 
the  composition  prescribed.  The  amount  can  be  learned  most  easily 
by  the  following  arbitrary  method. 

CHARGE   SHEET. 


Name  of  Ore 

Weight 

SiO2 

Fe 

Bases 

A1203 

S 

H2O 

Wet 

Dry 

% 

Ib. 

* 

Ib. 

* 

Ib. 

% 

Ib. 

Ib. 

Masabl  No.  1552... 

12.0 

1120 

1000 

4.0 

40 

60.0 

600 

1.0 

10 

Maxwell  No.  27... 
Limestone 

6.0 
0.0 
3.0 

1060 
1236 

1000 
700 
1200 

18.0 
3.0 
4.4 

180 
21 

293 

50.0 
1  0 

500 
12 

1.0 
53.0 
1.0 

10 
371 
12 
393 

2.0 
0.8 
3.5 

20 
6 
42 

78 

1.0 

12 
12 

Coke  

1112 

Slag                                           For  Pig    =   64                                       330                            Pig 

SiO2     =34%                                       For  Slag  =  229                                            63                          Fe  =  94.0 
A1203  =  14                                                                                                                               Si02  =  5.4,  81  =    2.5 

g£  ;1}«-  «o,  x  1.44                                                o-j.4 

FeO      =0.8                                                                                                                                            8    =    1.5 

8     =    1.4 


For  an  approximation,  we  take  an  amount  (to  the  nearest 
hundred  pounds)  of  limestone  (700  Ib.)  equal  to  twice  the  silica 
and  alumina  of  the  ore.  The  weight  of  all  the  constituents  are 
carried  out,  summing  them  up  as  shown.  The  1200  Ib.  of  pig  is 
to  contain  5.4%  Si02,  or  64  Ib.,  and  this  deducted  from  the  total, 
leaves  229  Ib.  available  for  the  slag.  In  the  assumed  slag  the  ratio 
of  the  silica  to  the  bases  (CaO  +  MgO)  is  as  34  to  49,  or  1  to  1.44. 
For  the  229  Ib.  of  silica  we  therefore  need  229  X  1-44  =  330  Ib.  of 
CaO.  But  we  have  393  Ib.  of  lime  already,  or  an  excess  of  63  Ib. 
Since  limestone  contains  approximately  half  the  weight  CaO,  this 
becomes  126  Ib.  limestone  or,  neglecting  decimals  less  than  10  Ib., 
it  becomes  120  Ib.  Erase  the  700  Ib.  replacing  it  by  the  new 
quantity  580  Ib.,  and  amend  the  calculated  amounts  in  the  same 
line  and  columns,  within  the  limit  of  error  (10  Ib.)  in  weighing. 


OF    THE    COMMON    METALS.  293 

The  actual  percentage  of  alumina  will  be  in  the  proportion 
229  :78 :  :34.0%  :12.2%.  We  can  decrease  the  alumina  by  using 
less  Maxwell  ore ;  or  we  can  choose  from  the  following  table  of 
typical  slags,  the  one  nearest  in  alumina,  which  would  be  the  one 
marked  'Lake  ore'  containing  Si02  35.5%,  bases  49.4%,  and  A1203 
12.0%.  Then  a  new  calculation  could  be  made,  using  the  new  slag 
in  place  of  the  one  first  proposed.  Either  slag  should  work  well  in 
the  furnace. 

TABLE  OF  TYPICAL  IRON  SLAGS. 


Middlesborough,    England. 
Saulines    France 

Si02. 

.  .     30.0 
31  6 

FeO. 

0.7 
0  6 

CaO. 

32.8 
47  2 

MgO. 
5.3 
1  4 

AL03. 

28.0 
17.0 

CaS 
19 

Cuban  ore         .              .  .    .  . 

.     33  2 

40.7 

11.1 

13.7 

Lake  ore   

35.5 

40.5 

8.9 

12.0 

Lake  ore    

.  .     34.7 

1.3 

40.1 

10.9 

11.3 

Cuban  ore   . 

34.5 

46.5 

7.9 

10.5 

79.     PIG  IRON. 

The  iron  produced  in  the  blast-furnace  is  not  pure,  but  contains 
3%  to  4%  carbon  and  1%  to  3%  silicon.  Some  of  the  carbon  is 
combined  chemically,  some  separated  by  graphite.  If  a  large 
proportion  is  combined,  the  metal  is  hard  and  the  fracture  of  the 
iron  looks  white.  If  a  large  proportion  is  free,  the  fracture  is  a 
gray  or  black  with  scales  of  graphite,  and  the  iron  is  soft  and  tough. 

Under  'Chemical  Reactions  of  the  Iron  Blast-Furnace,'  equation 
3  indicates  the  formation  of  carbon,  and  equation  8  the  reduction 
of  silica  to  silicon,  both  elements  entering  the  pig-iron.  A  small 
amount  of  sulphur,  seldom  less  than  0.2  and  often  0.25%  or  more, 
is  present.  As  the  amount  increases  above  0.1%  the  iron  becomes 
harder  and  more  brittle. 

The  percentage  of  silicon  and  sulphur  in  the  iron  depends  in  large 
measure  upon  furnace-conditions;  hence  it  can  be  controlled;  but 
all  the  phosphorus  present  enters  the  pig-iron.  In  pig-iron  for  steel 
manufacture  by  the  usual,  or  acid  bessemer  process,  the  phosphorus 
in  the  pig  must  not  exceed  0.10%.  Therefore,  in  the  ore,  it  must 
not  be  higher  than  0.05  or  0.06%.  In  the  acid  bessemer  process  the 
phosphorus  is  not  eliminated,  and  it  tends  to  make  steel  red-short 
(brittle  when  hot)  a  quality  that  interferes  with  the  subsequent 
rolling.  Phosphorus,  on  the  other  hand,  imparts  the  quality  of 
fluidity  to  cast-iron.  Iron  that  contains  3%  P  is  in  demand  where 
intricate  castings  are  to  be  made,  and  can  be  used  where  brittleness 
is  of  minor  importance. 

Cast-iron  as  compared  with  steel  and  wrought  iron  has  the 
following  characteristics : 


294  THE    METALLURGY 

(1)  It   is   brittle   because   of   the   presence   of   the   metalloids, 
carbon,  and  silicon. 

(2)  Because  of  the  presence  of  the  metalloids  it  is  fusible;  and 
it  derives  thus  the  most  valuable  property.     It  runs  freely  from 
the  blast-furnace  and  can  be  cast  in  intricate  molds  to  form  castings 
of  any  kind.     Wrought   iron   at  the   same  temperature  would  be 
pasty  and  would  not  run.     Steel,  which  is  intermediate   between 
wrought-iron  and  cast-iron  in  the  contained  carbon,  can  be  made 
into  castings,  however,  but  not  readily  like  cast-iron.     The  making 
of  steel-castings  is  becoming  more  common. 

(3)  It  cannot  be  forged  either  hot  or  cold. 

Pig-iron  is  graded  according  to  the  appearance  or  the  fracture 
of  a  freshly  broken  piece,  into  several  classes  or  grades  ranging 
from  a  soft- gray  to  a  white  iron  as  below. 

ALABAMA    PIG    IRON. 

Graphite  Combined 

Grade.                                              carbon.  carbon.  Silicon. 

Silver  gray    3.13  0.02  5.5 

No.   1  soft 3.48  0.03  3.5 

No.   2  soft 3.53  0.03  3.5  to  4.0 

No.   1   foundry 3.49  0.07  2.8  to  3.5 

No.  2  foundry 3.55  0.07  2.2  to  2.6 

No.  3  foundry 3.48  0.10  2.0  to  2.4 

Gray  forge 3.00  0.57  1.3  to  1.7 

Mottled   2.11  1.22  1.1  to  1.6 

White    0.10  2.92  0.7  to  1.2 

This  table  shows  the  increase  of  combined  carbon,  and  the 
decrease  of  silicon,  as  the  grade  approaches  white  iron. 

The  first  grades  are  more  difficult  to  make,  and  command  a 
higher  price  as  shown  by  the  following  quotations  on  Southern  or 
Alabama  pig-iron,  per  long  ton : 

No.  1  soft    $18.75 

.  No.  2     "      18.25 

No.  1  foundry    18.75 

No.  2     "              18.25 

No.  3     "              17.50 

No.  4     "              17.00 

Gray  forge    16.50 


PART   VI.     COPPER 


PART  VI.  COPPER. 

80.  COPPER  ORES. 

We  are  to  think  of  copper  ores  as  mineral  aggregates,  carrying 
frequently  10  to  15%  copper  or  less,  with  associated  minerals  and 
an  earthy  gangue.  Treating  the  ore,  is  a  problem  not  only  of 
obtaining  the  copper,  but  of  separating  and  eliminating  the  gangue 
and  associated  minerals.  Though  there  are  many  kinds  of  copper 
ores,  those  of  commercial  importance  are  few  in  number.  We  may 
divide  them  into  three  classes  (1)  the  sulphides;  (2)  the  oxides, 
including  the  carbonate  and  silicate;  (3)  ores  containing  native 
copper. 

Sulphides. — Chalcopyrite,  CuFeS2,  when  pure  contains 
copper.  This  is  by  far  the  most  widely  distributed  and  most 
abundant  of  the  ores  of  copper,  and  furnishes  the  world's  principal 
supply  of  the  metal.  It  is  frequently  accompanied  by  iron  pyrite, 
and  has  silicious  gangue  even  when  the  sulphide  is  massive.  In 
consequence,  the  ore  often  carries  no  more  than  3  to  4%  copper, 
but  it  is  particularly  suited  to  pyrite  smelting.  Silver  and  gold  are 
found  in  the  ore  in  small  quantity.  The  deposits  at  Mt.  Lyell, 
Tasmania,  that  have  been  so  successfully  worked  by  pyritic  smelting, 
are  chiefly  of  massive  iron-pyrite,  containing  chalcopyrite,  and 
carrying  4.5  to  5%  copper  with  0.15  oz.  gold  and  3  oz.  silver  per 
ton. 

Chalcocite  (copper-glance),  Cu2S,  is  computed  to  contain  79.7% 
copper,  but  it  is  seldom  pure  even  in  the  crystalline  form,  the 
copper  having  been  replaced  by  iron  and  other  metals.  The  impure 
mineral  shows  the  characteristics  of  the  pure  mineral  when  carrying 
as  little  as  55%  copper.  The  pure  crystals  resemble  the  artificial 
product  'white  metal,'  a  high-grade  copper  matte  produced  in  the 
furnace. 

Bornite  (peacock  ore),  Cu3FeS3,  when  pure  contains  55.6% 
copper.  It  is  found  associated  with  chalcopyrite  and  chalcocite  in 
proportions  varying  from  42  to  70%  copper,  without  losing  the 
characteristic  varied  colors. 


298  THE    METALLURGY 

; 

Enargite  (4CuS  +  Cu2SAs2S3),  48.6%  copper,  an  arsenide, 
occurs  in  Butte,  Montana,  ores. 

Tetrahedrite  (gray-copper,  fahlerz),  (Cu2S,FeS,ZnS,Ag2S,PbS), 
(Sb2S3,As2S3),  may  be  computed  as  containing  30 .4%  copper,  but 
it  varies  greatly  in  the  copper  and  silver  content.  It  has  already 
been  mentioned  as  a  silver  ore.  Because  of  the  contained  arsenic 
and  antimony,  it  is  unfavorable  as  a  copper  ore,  and  it  is  only  because 
of  the  richness  in  silver  that  it  is  treated. 

Oxides,  carbonates,  and  silicates. — These  ores  are  the  result  of  the 
decomposition  of  the  copper  sulphides,  by  air  and  water.  We  find 
them  in  the  upper  zones  of  mineral  deposits  accompanied  by  iron 
oxide  which  also  is  the  result  of  the  decomposition  of  iron  sulphide. 
As  we  sink  on  the  vein  we  find  the  oxidized  ore  of  the  upper  levels 
giving  place  in  depth  to  the  unaltered  sulphide. 

Cuprite  (red  copper-oxide),  Cu2O,  88.8%  copper,  is  a  product 
of  decomposition.  It  often  permeates  large  masses  of  iron  ore. 
Large  lumps  of  the  ore  are  sometimes  found,  the  center  of  which 
contains  unaltered  metal.  These  evidently  are  the  result  of  the 
oxidation  of  a  mass  of  native  copper. 

Melaconite  (black  oxide  of  copper),  CuO,  contains  when  pure 
79.8%  copper.  The  ore,  with  the  copper  in  part  replaced  by  oxides 
of  iron  and  manganese,  is  sometimes  found  in  masses  large  enough 
to  pay  for  extraction,  and  containing  20  to  50%  copper.  The  so- 
called  black  oxide  of  the  Blue  Ridge  region,  on  the  border  of 
Tennessee,  North  Carolina,  and  Virginia,  seems  to  be  an  intimate 
mixture  of  copper  glance,  black  copper-oxide,  copper  carbonate,  and 
native  copper  with  iron  oxide  and  sulphide.  The  ore  can  be  readily 
roasted  in  lump  form. 

Malachite,  CuCO3,  CuOH20,  57.3%  copper,  occurs  widely 
distributed,  ordinarily  in  non-paying  quantities  as  a  decomposition 
product  in  surface  deposits,  but  sometimes  sufficiently  rich  to  work. 
It  is  found  mixed  with  limestone,  dolomite,  oxides  of  iron  and 
manganese,  and  silica.  It  is  difficult  to  judge  the  copper-content  of 
the  ore  from  the  appearance,  but  the  green  color  makes  its  presence 
readily  recognizable. 

Azurite,  2CuCO3  -f  Cu2OH2O,  is  computed  to  contain  55.2% 
copper.  The  ore  is  blue,  as  the  name  indicates,  and  the  appearance 
is  striking.  It  occurs  in  the  same  way  as  malachite,  and  often  is 
associated  with  malachite,  but  it  is  less  abundant,  and  often  is  only 
a  coloring  on  other  oxides. 

Chrysocolla,  a  hydrated  silicate  of  copper,  containing  when  pure 


OF    THE    COMMON    METALS.  299 

40%  copper,  is  a  decomposition  product  of  copper  sulphide,  and  is 
often  accompanied  by  malachite. 

Native  copper. — Native  copper  is  found  extensively  in  the  copper 
region  of  Lake  Superior.  Elsewhere  it  occurs  sparingly  and  is 
not  commercially  important  though  it  often  accompanies  the  oxidized 
ores.  In  the  Lake  Superior  region  it  is  found  in  wide  lodes 
disseminated  through  the  lode-matter  0.5%  to  4%  of  the  whole,  and 
eveji  when  of  the  lowest  grade  mentioned,  by  concentrating  can  be 
recovered  at  a  profit.  The  concentrate  or  'mineral/  as  it  is  locally 
named,  is  produced  in  different  grades,  ranging  from  30  to  94% 
copper.  Much  of  the  native  copper  is  pure ;  in  other  instances  it 
carries  a  little  arsenic. 

81.     THE  EXTRACTION  OF  COPPER. 

Copper  may  be  extracted  from  the  ore  by  dry  or  by  wet  methods. 
By  the  dry  method  the  ore  is  smelted,  the  process  being  one  of 
igneous  fusion.  By  the  wet  or  hydro-metallurgical  methods  the 
copper  is  leached  from  the  ore.  The  striking  point  of  difference 
between  the  two  methods  is  that,  in  the  first,  we  change  the  form 
of  the  entire  mass  by  smelting  into  a  fluid,  effecting  then  a  separation 
of  the  copper  from  the  worthless  part,  while  in  the  wret  method 
we  act  upon  the  copper  alone,  leaving  the  greater  part  of  the  ore 
in  the  original  condition. 

The  dry  or  igneous  methods. — Probably  more  than  90%  of  the 
world's  production  of  copper  is  by  smelting.  The  methods  of 
smelting  vary  with  the  nature  of  the  ore.  We  may  divide  them  into 
the  following: 

(1)  The    smelting    of   the    oxidized    ores    of   copper    in    blast- 
furnaces. 

(2)  The  blast-furnace  matte-smelting  of  the  roasted  sulphide  of 
copper. 

(3)  The  blast-furnace  matte-smelting  of  raw  or  unroasted  copper 
sulphides   (pyrite  matte-smelting). 

(4)  The   reverberatory   smelting   of  roasted   sulphide   with   or 
without  oxidized  ore. 

Of  the  above  methods,  the  first  gives  the  copper  in  metallic 
form,  while  the  second  and  third  are  merely  concentration  processes 
by  which  the  copper  is  concentrated  into  a  small  product  from  a 
large  bulk  of  the  original  ore.  The  product,  or  matte,  has  to  be 
further  treated  before  the  copper  is  recovered  from  it.  The  copper, 
or  copper-matte,  produced  in  the  above  operations  carries  with  it 


300  THE    METALLURGY 

the  gold  and  silver  of  the  ore.     It  thus  becomes  a  collector  of  the 
precious  metal. 

82.     COPPER  SMELTING  OF  OXIDIZED  ORES. 

This  resembles  the  smelting  of  iron  ores  in  that  metal  is  obtained 
in  metallic  form  in  one  operation.  The  ore,  which  does  not  contain 
sulphur,  is  charged  into  a  blast-furnace  and  smelted  with  coke  for 
fuel.  The  products  are  slag  and  blister-copper,  the  latter  being 
metallic  copper  containing  impurities  that  have  been  taken  up 
in  smelting,  much  as  carbon  and  silicon  are  absorbed  in  iron  smelting. 

In  the  southwestern  part  of  the  United  States  (New  Mexico  and 
Arizona)  and  in  Mexico,  ores  are  found  that  contain  copper  in  the 
form  of  malachite,  cuprite,  chrysocolla,  and  the  native  metal. 
Being  nearly  free  from  copper  sulphide,  they  have  been  treated  for 
the  recovery  of  the  copper  by  a  single  operation.  As  compared  with 
matte  smelting  the  process  has  the  advantage  in  that  it  yields  nearly 
pure  copper  (97  to  98%)  while  in  matte-smelting  the  blast-furnace 
product  is  40  to  50%  copper.  The  objection  to  the  method  has  been 
that  separation  of  copper  from  the  slag  is  not  complete.  The  slag 
resulting  from  the  operation  is  1.5  to  2%  copper,  while  in  matte- 
smelting  it  is  only  0.5%.  The  method  therefore  has  fallen  into 
disuse,  and  can  be  revived  only  where  suitable  oxidized  ore  is  found, 
distant  from  railroads,  and  where  no  sulphides  are  present  to  mix 
with  it  and  form  matte. 

Fig.  128  is  a  sectional  elevation  of  a  blast-furnace  building 
suited  to  the  smelting  of  oxidized  copper  ores.  It  has  two  floors 
or  levels,  the  upper,  called  the  charge-floor,  and  the  lower,  the  slag- 
floor.  The  ground  at  the  right  drops  away  and  furnishes  a  place 
for  a  dump.  The  ores  are  stored  in  bins  at  the  charge-floor  level, 
and  are  brought  in  weighed  charges  to  the  furnace  door,  the  sill  of 
which  is  flush  with  the  charge-floor.  This  door  is  shown  also  in 
Fig.  129.  The  slag  and  blister  copper  are  withdrawn  near  the  bottom. 
Behind  the  furnace  is  seen  the  pipe  or  blast-main  by  which  air  is 
conducted  to  the  wind-box  at  the  tuyeres.  The  furnace  stack  extends 
above  the  roof,  and  removes  the  escaping  gas. 

Fig.  129  is  a  view  of  a  blast-furnace  for  the  production  of  blister 
copper  from  oxidized  copper  ores.  At  the  bottom  are  seen  the  two 
half  drop-doors  which,  when  the  furnace  is  running,  are  swung  up 
into  position  to  close  the  bottom  of  the  furnace.  On  them  is  placed 
the  bottom  lining  of  brick.  The  wall  within  the  curb  is  lined  with 
fire-brick  as  high  as  the  tuyeres  forming  the  crucible.  The  water- 
jacket,  including  the  tuyere-openings  into  the  furnace,  extends  up 


OF   THE    COMMON    METALS. 


301 


302  THE    METALLURGY 

to  the  feed  floor.  It  consists  of  an  inner  and  an  outer  plate  with 
space  between  in  which  water  flows,  continually  escaping  through 
outlet  pipes  at  the  top  of  the  jacket.  Thus,  while  the  inner  plate 
is  in  contact  with  the  highly  heated  content  of  the  furnace,  it  is  kept 
cool  and  is  not  melted  or  attacked  by  the  slag.  The  water-jacket 
is  surrounded  by  a  wind-box  as  shown.  There  are  six  tuyere- 
openings,  having  covers  with  mica-covered  peep-openings  by  means 
of  which  the  condition  of  the-  tuyeres  can  be  observed.  The  base- 
plate rests  upon  four  iron  columns  that  are  supported  by  a  solid- 
stone  or  concrete  foundation. 

The  crucible  contains  the  molten  contents  of  the  furnace,  the 
copper  below  and  the  lighter  slag  floating  upon  it.  There  are  two 
tap-holes  and  two  spouts  ;  the  lower,  close  to  the  bottom,  is  to 
remove  the  molten  copper,  the  upper,  a  few  inches  higher,  is  to 
withdraw  the  slag.  From  time  to  time,  as  slag  or  copper  accumulates, 
it  is  withdrawn  by  piercing  a  hole  through  the  clay-stopping  of  the 
tap-holes  by  means  of  a  pointed  steel  tapping-bar.  The  flow  is 
arrested  by  thrusting  into  the  opening  a  plug  of  clay  stuck  on  the 
end  of  a  button-headed  stopper-rod  or  dolly.  The  slag  is  received 
into  a  slag-pot,  Fig.  135,  mounted  on  wheels,  which  when  filled,  is 
pushed  out  and  poured  over  the  edge  of  the  dump.  The  copper  is 
received  in  a  'bullion  mold,'  which  after  filling,  is  set  aside  until 
the  copper  has  solidified.  The  ingot  is  then  dumped  out  and  the 
mold  again  used.  The  furnace  shown  is  a  round  one,  36  in.  diam. 
inside,  thus  having  a  bosh  or  enlargement  of  six  inches  on  the  side. 

In  operation,  the  furnace  is  kept  full  to  the  feed-door  with 
alternate  layers  of  fuel  and  charge.  The  blast  rises  through  the 
column  of  materials  a  distance  of  seven  feet  and  passes  off  through 
the  stack,  which  has  sufficient  draft  to  take  away  the  gas  and  smoke 
and  also  the  air  that  enters  the  feed-opening  or  door.  The  blast 
enters  under  a  pressure  of  12  oz.  per  sq.  in.  causing  an  intense 
combustion  of  the  coke  and  the  fusion  of  the  charge.  The  copper, 
i  educed  by  the  glowing  coke,  collects  in  drops  and  finds  its  way 
to  the  bottom  of  the  crucible,  while  the  gangue  of  the  ore,  fluxed  by 
the  addition  of  iron  ore  and  limestone,  forms  a  fusible  slag. 

The  cost  of  treatment  of  oxidized  copper  ore  may  be  stated  to 
be  $8  per  ton.  The  slag  contains  1.5  to  2.5%  copper,  so  that,  under 
the  best  circumstances,  we  rely  on  recovering  82%  copper,  while 
of  the  silver  91  is  obtained. 


83.     MATTE-SMELTING. 
If  we   smelt   copper-bearing   sulphide   ore   in   the   blast-furnace 


OF   THE    COMMON    METALS.  303 

described,  with  added  fluxes  to  form  a  fusible  slag,  we  form  an 
artificial  sulphide  or  matte  of  23  to  25%  sulphur.     The  copper  that 


Fig.    129.      BLAST-FURNACE    FOR  OXIDIZED   COPPER  ORES. 


304  THE    METALLURGY 

was  in  the  ore  and  the  iron  from  the  charge  enter  the  matte,  but 
the  quantity  of  matte  so  formed  is  but  little  less  than  that  of  ore 
originally  put  into  the  furnace.  If,  however,  we  first  roast  the 
ore,  the  quantity  of  sulphur  present,  and  consequently  the  amount 
of  matte  made,  is  less,  and  the  ratio  of  the  ore  to  the  matte  may 
be  five  or  ten  to  one.  In  the  matte  will  be  the  copper  as  a 
sulphide,  and  in  forming,  the  matte  will  take  up  the  silver  and 
gold  of  the  ore.  By  this  operation  we  collect  the  precious  metals 
in  a  product  one-fifth  to  one-tenth  the  original  ore,  the  matte 
being  termed  a  'collector.'  It  then  can  be  further  treated  to  convert 
it  into  metallic  copper  carrying  the  precious  metals.  By  the  process 
of  electrolytic  refining,  the  gold,  and  silver  are  eventually  separated 
from  the  copper.  A  charge,  suited  to  matte-smelting,  methods, 
would  therefore  consist  of  roasted  ore  retaining  1%  sulphur  with 
oxidized  copper  and  copper-free  ores  containing  gold  and  silver 
added  for  the  precious-metal  content.  It  would  also  carry  fluxes 
to  make  a  fusible  slag  and  supply  iron  for  the  matte. 

The  products  of  the  furnaces  are  slag  and  matte.  The  former 
is  the  result  of  the  union  of  the  silica  in  the  charge  with  the 
various  bases,  chiefly  iron  oxide  and  lime.  The  latter  is  the  complex 
artificial  sulphide  produced  by  the  sulphur  in  the  charge  combining 
with  copper  and  iron.  The  affinity  of  sulphur  for  copper  is  greater 
than  for  iron,  and  it  takes  the  former  first;  then  if  it  needs  iron 
it  takes  that  also,  until  a  compound  of  both  has  been  formed 
that  contains  approximately  25%  sulphur.  Any  further  iron 
present  enters  the  slag.  As  the  ratio  of  the  amount  of  the  iron 
thus  available,  to  the  amount  of  other  bases,  varies,  so  will  the 
slag  vary  in  composition;  but  the  principal  requirement  is  that 
the  quantity  of  silica  be  enough  to  form  a  fusible  slag.  Slags  of 
25  to  40%  silica  are  common  in  copper-matting  practice,  and  the 
lower  limit  is  sometimes  exceeded,  when  for  economy  in  smelting, 
it  is  desired  to  use  as  little  flux  as  possible. 

84.     THE  COPPER-MATTING  BLAST-FURNACE. 

For  producing  matte  from  copper-bearing  ores,  whether  these 
are  to  be  smelted  after  roasting  or  treated  raw  by  the  method  of 
pyritic  smelting,  we  use  the  furnace,  Fig.  129,  already  described, 
or  one  of  the  rectangular  type,  Fig.  130  and  131. 

Fig.  130,  at  the  left,  represents  a  transverse  sectional  elevation 
of  a  furnace  of  42  by  120  in.  interior  hearth-dimensions,  having  18 
tuyeres,  nine  at  each  side,  and  a  capacity  of  150  tons  of  charge 
daily.  Fig.  131  is  a  perspective  view  of  a  similar  furnace,  showing 


OF    THE    COMMON    METALS. 


305 


the  iron-work  as  high  as  the  feed-floor,  but  differing  from  Fig.  130 
in  having  a  trapped  slag-spout,  more  fully  shown  in  Fig.  133. 

The   sole-plate   of  the   furnace   rests   on   jack-screws,    and   can 
be  lowered  and  set  aside  when  it  is  desired  to  make  repairs  or 


i  i     ! 

JbmJh. 

Fig.    130.      COPPER  MATTING  BLAST-FURNACE. 

to  clean  out  the  furnace.  It  is  protected  from  the  action  of  the 
molten  matte  and  slag  by  a  9-in.  layer  of  fire-brick.  In  Fig.  130 
are  seen  crucible  plates  which  rest  upon  the  sole-plate.  These 
are  lined  with  18  in.  of  brick.  The  jackets  shown  in  Fig.  131 
extend  down  to  the  sole-plate  and  the  water-cooling  is  sufficient 
protection  from  the  action  of  the  molten  materials.  The  sole-plate 
within  the  furnace,  however,  is  covered  by  the  brick  lining.  The 


306 


THE    METALLURGY 


jackets,  shown  best  in  Fig.  132,  are  at  least  9  ft.  high,  and  in  the 
furnace  represented,  there  are  two  of  them  on  each  side,  and  one 
at  each  end.  At  one  end  the  jacket  is  shorter,  and  the  space 
below  is  filled  with  a  water-cooled  tap- jacket  through  which  the 
slag  is  withdrawn.  In  Fig.  130,  the  longitudinal  view  shows  the 
arrangement  of  a  furnace  with  three  jackets  at  each  side,  and 
two  jackets  at  each  end.  The  small  jackets  are  easily  handled  and 


Fig-.    131.      VIEW  OF  COPPER  MATTING    BLAST-FURNACE. 


replaced.  In  Fig.  I3&the  inlets  for  water  are  at  half  the  height 
of  the  bosh,  and  the  water  outlets  are  at  highest  point  to  keep  them 
full  of  water.  They  are  tied  or  clamped  together  with  heavy 
angles,  but  in  the  other  figures,  -with  I-beams. 

Fig.  133  is  a  water-cooled  trapped  spout  used  in  connection 
with  the  furnace  as  indicated  in  Fig.  131.  Through  it  now  the 
slag  and  matte.  Before  the  slag  can  overflow  it  must  fill  the  spout 


OF    THE    COMMON    METALS. 


307 


and  cover  the  outlet  or  tap-hole  through  the  jacket.  It  thus 
prevents  the  escape  of  the  blast,  and  flows  in  a  regular  stream  as 
fast  as  it  forms  within  the  furnace.  The  furnace  shown  in  Fig.  130, 


308 


THE    METALLURGY 


is  arranged  differently.  There  is  a  spout  and  a  water-cooled  tap- 
jacket  at  the  end  of  the  furnace  through  which  the  slag  is  removed, 
while  the  matte,  as  it  accumulates,  is  removed  by  a  spout  and  a 
side  tap-hole  at  the  level  of  the  crucible-bottom.  In  this  case,  the 
separation  between  the  slag  and  matte  is  effected  within  the 
furnace ;  in  the  former  case,  where  the  trapped  spout  is  used  (since 
matte  and  slag  issue  together)  they  are  separated  outside  the 
furnace  in  the  fore-hearth  or  settler  Fig.  134. 

The   fore-hearth,    made    of   cast-iron   plates,   4   by   6   ft.    inside 
dimensions,  is  lined  with  a  layer  of  brick,  and  is  mounted  on  wheels 


Fig.    133.      TRAPPED   SPOUT. 


so  that  it  can  be  quickly  set  aside  and  a  new  one  put  in  the  place 
when  needed.  The  slag  and  matte  flow  into  it  at  one  end  and  keep 
it  full  of  molten  slag.  At  the  other  end,  the  slag  flows  out.  The 
matte  settles  on  the  way  and  collects  in  the  bottom  of  the  fore- 
hearth,  and,  when  accumulated,  is  tapped  at  the  tap-hole  and 
spout,  seen  at  the  side.  Meanwhile  the  slag,  flowing  from  the  fore- 
hearth,  is  caught  in  slag-pots,  Fig.  135,  and  taken  to  the  edge  of 
the  dump  and  poured.  The  slag  cools  on  the  surface  of  the  fore- 
hearth  and  forms  a  crust  from  beneath  which  the  molten  slag 
flows.  Crust  forms  also  at  the  sides  and  bottom,  and  becomes 
gradually  thicker;  and  after  several  days  becomes  so  thick  that 
the  molten  part  of  the  interior  is  too  small  to  permit  of  a  good 


OF    THE    COMMON"    METALS. 


309 


separation  of  the  matte  from  slag.  When  this  results,  the  fore- 
hearth  is  pried  back  on  the  wheels,  and  replaced  by  another.  In 
the  furnace,  Fig.  131,  at  the  middle  side-jacket,  another  tap-hole 
furnished  with  a  spout  is  seen.  This  is  generally  kept  closed,  but 
is  opened  when  it  is  desired  to  empty  the  hearth  of  the  matte  and 
slag. 

The  transverse   view,   Fig.   130,   shows  the  side-jackets.     These 


Fig.    134.      FORE-HEARTH. 


have  brackets  or  knees  riveted  to  them  and  rest  on  I-beams  that 
are  secured  to  the  columns.  Thus,  when  the  sole-plate  is  removed, 
the  jackets  are  not  disturbed.  The  distance  between  the  side  jackets 
is  42  in.  at  the  tuyeres,  and  66  in.  at  the  top.  The  bosh,  or 
enlargement,  is  thus  12  in.  on  the  side.  Above  the  jackets  are  the 
cast-iron  distributing  plates,  forming  the  sills  of  the  feed-doors. 
The  feed-doors  in  the  opposite  long  sides  of  the  furnace  make  it 
accessible  from  end  to  end,  not  only  for  feeding  and  trimming  the 


310 


THE    METALLURGY 


charge,  but  for  cutting  with  chisel-bars  the  accretion  or  scaffolding 
that  may  form  on  the  interior  surface  of  the  jackets. 

The   portion  of  the   furnace   above   the   feed-floor   level,   called 


OF    THE    COMMON    METALS.  311 

the  stack  or  top,  is  of  brick  supported  by  a  deck-plate  or  mantel- 
plate  of  I-beams  resting  on  the  cast-iron  columns  that  extend  down 
into  the  foundation.  The  upper  portion  of  the  stack  is  a  hood  of 
sheet-steel  terminating  in  a  pipe  that  extends  through  the  roof  of 
the  furnace  building.  Sometimes  a  branch  pipe  leads  from  the  hood 
to  a  dust-chamber  where  the  dust  is  collected. 

The  bustle-pipe,  by  which  the  blast  at  a  pressure  of  %  to  2  Ib. 
is  brought  to  the  furnace,  extends  around  three  sides  and  connects 
by  the  sheet-metal  pipes  to  the  tuyeres,  shown  in  detail  in  Fig.  136. 
The  tuyere  is  6  in.  diam.  and  has  a  6-in.  screw-cap  into  which  is 
inserted  a  nipple  with  a  cap  having  a  mica-covered  peep-hole, 


Fig.    136.      TUYERE. 

through  which  the  condition  of  the  furnace  can  be  observed.  At  the 
branch  above  the  tuyere  is  shown  a  slide-valve.  In  Fig.  131,  just 
below  the  bustle-pipe,  is  a  waste  launder  to  receive  the  overflow 
from  the  jackets,  and  below  it  is  a  3-in.  water-supply  pipe  branching 
to  each  jacket,  and  to  the  water-cooled  trapped  spout  at  the  front. 

85.     LARGE  FURNACES  AND  FORE-HEARTHS. 

The  tendency  of  late  years  has  been  to  increase  the  size  of 
copper-matting  blast-furnaces.  Increase  in  width  would  require 
higher  blast-pressure  to  drive  the  air  to  the  center  of  the  furnace ; 
hence  increased  capacity  had  to  be  gained  by  increasing  the  length 
of  the  furnace.  At  the  same  time,  to  supply  more  air,  the  blast 
presure  has  been  increased  in  some  cases  to  40  oz.  or  21/2  Ib.  per 
sq.  in.  A  furnace  56  by  180  in.  under  these  conditions  smelts  400 
tons  of  ore  daily.  For  the  matte  to  properly  settle  from  the  slag, 


312  THE    METALLURGY 

with  so  large  a  flow,  a  cylindrical  fore-hearth  has  been  used  16  ft. 
diam.  by  5  ft.  deep,  exterior  dimensions,  lined  with  fire-brick.  In 
the  pool  of  molten  material  inerasted  with  slag,  the  separation  is 
effected.  The  molten  content  of  the  furnace  flows  into  it  on  one 
side  and  the  slag  flows  out  at  the  opposite  side.  From  time  to 
time  the  fore-hearth  is  tapped  at  the  lower  tap-hole,  and  five  to 
ten  tons  of  matte  are  drawn  into  a  ladle,  that  takes  the  matte 
to  a  converter  for  further  treatment  to  obtain  the  copper  contained. 
The  lengthening  of  the  furnace  has  been  carried  so  far  that,  at 
the  Washoe  plant,  Anaconda,  Montana,  a  furnace  51  ft.  long,  having 
1600  tons  daily  capacity  has  been  for  some  time  in  operation,  and 
recently  one  87  ft.  long  and  3000  tons  daily  capacity  has  been  built 
and  operated.  The  first  furnace  has  two  fore-hearths,  each  16  ft. 
diam.,  and  the  second  three  of  that  size. 

Ore  for  matte-smelting. — Ores  suitable  for  matte  smelting  are 
the  oxidized  ones  containing  little  or  no  sulphur ;  ores  that  have 
been  roasted;  and  ores  that  are  silicious  and  oxidized  containing 
gold  and  silver.  Ores  that  require  to  be  first  roasted  are  better 
roasted  in  lump  form  in  heaps  or  stalls,  for  the  reason  that  'the 
blast-furnace,  because  of  its  strong  air  currents,  is  not  suited  to 
smelting  fine  ore.  The  amount  of  matte  made  (matte  fall)  depends 
upon  the  quantity  of  sulphur  in  the  charge,  and  to  get  sufficient 
concentration  (little  matte  from  much  ore)  the  sulphur  is  kept 
low.  To  the  ore  above  described  is  added  limestone  and  iron  ore 
as  flux,  and  a  quantity  of  fuel  equal  to  10  to  15%  of  the  charge. 

Starting  the  furnace. — Since  a  blast-furnace  is  water-jacketed 
the  operation  of  warming  it  is  a  simple  one.  The  fire-brick  lining 
of  the  crucible  is  dried  and  warmed  by  several  hours  heating  with 
a  wood  fire.  The  end  and  side  tap-jackets  are  removed  to  permit 
the  air  to  enter  to  the  fuel.  When  the  hearth  is  hot  the  wood  ashes 
are  scraped  out  and  a  fresh  fire  of  wood,  filling  the  crucible  a 
foot  deep,  is  started,  wood  of  uniform-sized  pieces  for  uniform 
burning  being  selected.  Upon  the  wood  is  placed  charcoal,  and 
upon  the  charcoal,  coke,  until  the  surface  is  1%  to  2  ft.  above  the 
tuyeres.  The  fire  is  increased  uniformly  and  regulated  by  checking 
the  draft  at  the  front  and  admitting  air  at  the  rear  as  required. 
When  the  coke  is  thoroughly  ignited  the  furnace  is  ready  for 
charging. 

The  brick-lined  fore-hearth  is  warmed  while  warming  the 
furnace.  The  wood  is  placed  carefully  against  the  walls,  leaving 
the  center  clear  for  the  air  to  reach  the  fuel  so  that  the  burning 
may  proceed  actively.  As  the  wood  burns,  charcoal  and  ashes 


OF    THE    COMMON    METALS.  313 

accumulate,  and  are  shoveled  out,  since  otherwise  they  form  a  layer 
through  which  the  heat  does  not  penetrate. 

Suppose  the  charge  of  ore  and  flux  to  be  2000  Ib.,  and  that  we 
intend,  as  in  regular  work,  to  use  with  it  12%  coke  or  240  Ib.  per 
charge.  We  put  in  a  layer  of  240  Ib.  coke,  then  one  of  500  Ib. 
slag.  This  is  followed  by  a  half  dozen  charges  each  of  240  Ib. 
coke  alternated  with  1000  Ib.  slag.  Next  we  put  in  a  half  dozen 
charges  of  240  Ib.  fuel  and  2000  Ib.  slag,  so  that  the  slag  when 
melted  shall  entirely  fill  the  fore-hearth.  We  now  begin  feeding 
the  regularly  calculated  charges  and  required  fuel.  At  this  time 
the  blast  is  admitted,  gently  at  first,  gradually  increasing  during 
a  half  hour,  after  which  the  furnace  should  be  in  full  blast.  Extra 
men  should  now  assist  in  charging  to  rapidly  fill  the  furnace.  At 
the  slag  floor,  before  the  blast  is  turned  on,  the  tap-jackets  and 
the  trapped  spout  are  put  in  place,  and  all  openings  closed  with 
plugging  mixture  or  'adobe.'  The  adobe  may  be  obtained  from 
a  neighboring  bank,  if  of  suitable  quality,  or  may  be  made  from 
coarsely  ground  fire-brick  mixed  with  clay. 

As  the  smelting  proceeds,  by  looking  into  the  tuyeres,  we  see 
that  the  slag  is  rising  to  their  level.  We  then  open  the  tap-hole 
and  permit  the  slag  to  flow  into  and  quickly  fill  the  fore-hearth. 
The  excess  steadily  overflows  to  the  slag-pots  set  to  catch  it,  or 
it  may  be  granulated  and  removed  by  water. 

When  feeding  the  furnace  care  is  taken  to  distribute  the  charge 
evenly,  not  feeding  coarse  ore  in  one  place  and  fine  in  another. 
Unless  we  exercise  care  in  this  regard  we  have  irregular  operation, 
blast  and  flame  coming  up  in  one  place  and  the  charge  looking  dead 
in  another.  When  this  begins  to  occur  we  load  the  active  places 
with  charge,  feed  lightly,  and  use  coarser  material  where  there  is 
little  action. 

We  may  adopt  either  the  intermittent  or  the  continuous  method 
of  removing  the  slag  from  the  furnace.  In  the  intermittent  method, 
as  arranged  in  Fig.  130,  the  slag  is  tapped  from  time  to  time  as 
it  accumulates  as  already  described,  taking  care  that  it  does  not 
gather  in  Such  quantity  as  to  rise  to  and  run  into  the  tuyeres, 
or  to  'slag'  them  as  it  is  called.  By  the  continuous  method,  the 
slag  and  matte  flow  continuously  from  the  furnace  through  the 
trapped  spout,  Fig.  131  and  133.  The  molten  products  enter  a  fore- 
hearth  and  the  separation  of  slag  from  matte  is  there  made. 

86.     REGULAR   OPERATION   OF   THE   FURNACE. 

The  furnace  being  filled  to  the  feed-doors,  we  have  a  7-ft. 
smelting  column  (distance  from  tuyeres  to.  feed-door).  As  the 


314  THE    METALLURGY 

charge  smelts  and  the  molten  materials  are  withdrawn,  the  surface 
gradually  sinks,  making  room  for  further  additions.  The  coke 
is  first  added  in  a  layer  over  the  surface,  and  upon  it  is  spread  the 
weighed  charge.  The  air  under  pressure  from  the  blowers,  driven 
into  the  furnace  at  a  pressure  not  less  that  %  Ib.  or  12  oz.  per 
sq.  in.,  burns  the  descending  coke  mostly  at  the  tuyeres.  The 
resulting  gas  with  the  sulphur  dioxide  from  the  burning  sulphur 
in  the  charge  appears  as  a  whitish  smoke  mingled  with  dust 
mechanically  carried.  This  passes  from  the  furnace-top  directly 
into  the  air,  or  to  a  dust-chamber  and  thence  to  the  stack.  The 
sulphur  remaining  unites  with  the  copper  and  a  part  of  the  iron 
and  forms  a  matte  or  copper-iron  sulphide.  The  matte,  in  forming, 
comes  into  contact  with  the  gold  and  silver  contained  in  the  ore 
of  the  charge  and  absorbs  them.  The  coke  reduces  the  iron  not 
needed  for  the  formation  of  the  matte  to  ferrous  form,  and  the 
ferrous  iron,  with  the  lime,  alumina,  and  other  bases,  combines  with 
the  silica  to  form  a  slag,  fluid  at  the  high  temperature  prevailing 
at  the  tuyeres. 

The  molten  slag  and  matte  flow  from  the  furnace  to  the  fore- 
hearth  where  the  separation  is  effected,  and  the  supernatant  slag, 
freed  from  the  matte,  escapes  by  an  overflow  spout,  and  is  received 
into  slag-pots  and  conveyed  to  the  dump.  Another  means  of 
disposing  of  the  slag  is  to  allow  the  stream  of  slag  from  the  fore- 
hearth  to  fall  into  a  launder  and  be  caught  by  a  horizontal  jet 
of  water  to  break  it  into  drops  and  cool  it  in  granules  about  wheat- 
size.  The  granules  are  carried  away  by  the  water,  in  a  cast-iron- 
lined  launder  to  the  dump.  The  matte  is  tapped  from  the  fore- 
hearth  as  it  accumulates,  through  a  tap-hole  near  the  bottom,  and 
flows  over  the  matte  spout  shown  at  the  right  of  the  transverse 
section,  Fig.  130. 

87.     MATTE. 

Matte  is  an  artificial  sulphide  formed  in  smelting  in  result 
of  the  union  of  sulphur  with  bases.  Iron  sulphide  (FeS),  such  as 
is  used  in  the  making  of  hydrogen  sulphide  in  the  laboratory,  is 
the  simplest  form.  To  produce  it  in  small  quantities,  a  covered 
assay  crucible  may  be  filled  with  shingle-nails  and  brought  to 
a  white  heat  in  a  wind-furnace  and  roll-sulphur  added  gradually 
until  the  content  fuses.  The  sulphide  is  then  poured,  and  broken 
for  use.  In  smelting  a  charge  containing  sulphur,  scrap-iron  will 
take  up  the  sulphur  and  form  matte.  If  copper  oxide  or  copper 
sulphide  is  present  in  the  charge,  the  sulphur  takes  the  copper  to 
form  the  matte  in  preference  to  taking  the  iron.  When  the  copper 


OF   THE    COMMON    METALS.  315 

is  exhausted,  the  excess  of  sulphur  expends  any  combining  power 
that  may  remain  by  taking  iron.  We  thus  have  a  copper-iron 
sulphide,  called  copper  matte.  If  lead  or  nickel  are  present  in  the 
charge,  they  partly  enter  the  matte.  Magnetic  iron  oxide,  taken 
from  the  charge,  also  enters  the  matte.  Thus  we  get,  finally,  a 
complex  compound,  as  the  following  table  shows  : 

COMPOSITION  OF  COPPER  MATTE. 

Cu,  S,  Fe,  Fe3O4, 

%  %  %             %    Sp.  gr. 

Anaconda   reverberatory   furnace 60.76  23.25  11.43  1.13  5.4 

Reverberatory  furnace    (Butte,  Mont.) 29.41  23.70  25.35  12.60  4.8 

Blast-furnace     (Butte,    Mont.) 36.15  23.38  24.97  8.51  5.1 

Blast-furnace    (Jerome,   Ariz.) 55.00  23.96  13.85  2.58  5.3 

Blast-furnace    (Elizabeth,    Vt.) 21.36  22.95  41.03  10.44  4.7 

Blast-furnace    (Sudbury,    Canada) 24.54  23.24  28.65  7.32  5.1 

The  Sudbury  matte  contains  also  15.56%  nickel  replacing  copper. 
It  will  be  noted  that  the  percentage  of  sulphur  (23  to  24)  is 
approximately  the  same  in  all  cases. 

88.     SLAG. 

Variation  in  slag  composition  is  permissible  in  copper-smelting ; 
the  requirement  being  that  the  slag  be  fluid  to  flow  from  the  tap- 
hole  of  the  furnace.  Slags  having  the  maximum  content  of  silica 
and  of  bases,  as  shown  below,  are  employed  successfully  in  the 
blast-furnace. 

In  silver-lead  smelting  practice,  such  variations  are  not 
allowable.  Slags  varying  from  the  composition  found  in  practice 
to  be  satisfactory,  even  though  they  be  fluid  and  run  well,  carry 
off  both  lead  and  silver.  In  copper  practice  such  slags  would  be 
clean  and  free  from  copper. 

THE   COMPOSITION  OF   SLAGS. 

SiO2,        FeO,      A12O;1,      CaO,       MgO,     ZnO,        BaO, 

Minimum     20  2  2  2  0  0  0 

Maximum     57  70  18  40  8  20  42 

A  slag  low  in  silica  could  not  carry  much  of  the  alkaline-earth 
bases,  and  would  be  high  in  iron,  and  of  a  specific  gravity  over  3.7. 
A  silicious  slag  would  work  well  with  a  heavy  limey  base,  and 
would  have  a  specific  gravity  of  3.5.  Since  the  separation  of  slag 
from  matte  results  from  the  difference  in  specific  gravity  of  the 
two  substances,  we  expect  a  better  separation  the  lighter  and  more 
silicious  the  slags.  Matte  takes  up  zinc  sulphide,  where  much  is 
present,  and  becomes  lighter,  so  in  this  respect,  zinc  is  detrimental 
to  effective  separation. 


316 


THE    METALLURGY 


89.     CALCULATION    OF    CHARGE    FOR    MATTE    SMELTING. 

A  suitable  charge  may  consist  of  roasted  ore,  oxidized  copper- 
ore  and  silicious  ores  containing  gold  and  silver.  The  requirement 
is  that  the  charge  contain  copper-bearing  ore,  and  enough  sulphur 
to  form  with  the  copper  a  suitable  matte  that  will  take  up  the 
gold  and  silver  that  the  charge  contains.  Enough  flux  is  added 
to  the  charge  to  make  a  suitable  slag,  and  10  to  15%  coke  or  char- 
coal to  smelt  the  mixture. 

The  products  from  the  furnace  are  slag  and  matte,  the  former 
being  the  result  of  the  union  of  the  silica  of  the  charge  and  the 
fuel,  with  the  bases  that  are  present.  A  part  of  the  sulphur  in 
the  charge  is  volatilized  by  the  heat  of  the  furnace,  but  a  large 
part,  still  remaining,  combines  with  the  copper  and  a  part  of  the 
iron,  to  form  the  complex  sulphide  called  matte.  The  iron  not 
needed  for  the  matte,  enters  the  slag.  Since  the  copper-furnace 
slag  may  vary  within  wide  limits  we  use  a  silicious  ore  where 
silica  is  abundant  and  a  basic  one  where  plenty  of  iron  is  present, 
or  in  treating  basic  ores. 

CHARGE    SHEET.       REGULAR   MATTE    SMELTING. 


Name  of  Ore 

Weight 

Cu 

Si02 

Fe  and 
Mn 

CaO  and 
MgO 

S 

H2O 

Wet 

Dry 

% 

Wt. 

% 

Wt. 

% 

Wt. 

% 

52.0 
1.8 

S  ir 
Fell 

Wt. 

% 

Wt. 

Roasted  Ore 

3.0 
0 

1     0 

Cu<< 

Cui 
Cuir 
t  Fe  ' 
Fe  ' 

1000 
300 
130 

nSlag 
iMatl 

i      (t 

10.0 
-e  = 

100 

100 
4 

25.0 
4.2 

7.2 

Foi 
Fo 

250 
12 
9_ 
271 
r  Matt 
rSlag 

30.0 
1.2 

e  = 

Cx 

300 

2_ 

302 
129 
173 

i  and 

156 
2_ 

158 

i  Mat1 
iMat 

10.0 

,e  = 
te  = 

100 

100 
25_ 
75 
3_ 
225 

Limestone 

Coke  (10$) 

96 
225 

129 

Slag. 

Si02  =35% 

FeO  +  CaO  =  55% 

Other  bases  =  10% 

100% 


In  the  slag. 

FeO  =  173  X  |-  =  222 

CaO'  =  158 


Cu  +  Fe 


Actual    FeO  +  CaO  =  380 
Needed  —  425 


Matte. 

S  =  23% 

Cu  +  Fe     —  69% 

=  §-=  «   (factor) 


FeO  +CaO        55 


CaO  too  little 


=     45 


SiO2 


=  r^  =  1.57   (factor)  and  271  X  1.57  =  425  of  FeO  +  CaO. 


Above  is  given  a  charge  calculation,  in  which  the  problem  is 
to  treat  a  single  roasted  ore  producing  a  slag  of  a  pre-determined 


OF   THE    COMMON    METALS.  317 

composition  and  a  matte  that  will  take  up  the  copper  that  is  present. 
Limestone  is  the  only  flux  to  be  used.  The  charge  is  of  a  size  to 
fill  the  charge-car  or  buggy  in  which  it  is  brought  to  the  furnace. 

We  will  adopt  1000  Ib.  as  a  weight  of  the  roasted  ore,  having 
the  composition  Cu  10%,  SiO2  25%,  Fe  30%,  and  roasted  so  that 
10%  sulphur  remains.  This  is  to  be  smelted  with  limestone 
containing  Si02  4%  and  CaO  52%  to  produce  a  slag  of  Si02  35% 
and  bases  (FeO  and  CaO)  55%,  together  90%,  leaving  10%  to  allow 
for  other  elements.  The  slag  has  been  chosen  of  this  composition 
as  one  that  has  been  found  to  work  well.  The  coke  has  12%  ash 
that  consists  of  SiO,  60%,  Fe  10%,  and  CaO  15%.  These  figures, 
calculated  to  the  coke,  are  SiO2  7.2%,  F*e  1.2%,  and  CaO  1.8  per  cent. 

A  metallurgist,  accustomed  to  types  of  ore,  knows  with  some 
degree  of  precision  how  much  flux  he  needs.  Suppose  we  decide 
upon  300  Ib.  flux.  For  the  calculation  we  enter  on  the  charge- 
sheet  the  1000  Ib.  ore,  the  300  Ib.  limestone,  and  10%  of  these  or 
130  Ib.  coke  in  the  column  of  dry  weight.  When  the  exact  figures 
have  been  computed,  the  wet  weights  may  be  inserted  in  the 
adjoining  column,  using  the  figures  for  per  cent  given  in  the  column 
marked  H2O.  The  percentage  of  ore,  flux,  and  fuel  are  then  written 
in  the  appropriate  columns,  and  the  corresponding  weights, 
calculated  to  the  nearest  pound,  are  written  in  and  the  totals  added. 

Beneath,  and  at  the  left  of  the  sheet,  tabulate  the 
slag  composition.  Find  the  ratio  of  base  to  silica,  which  in  this 
case  will  be  1.57  to  1.  On  the  right  of  the  sheet  write  the  matte 
composition.  We  know  it  will  carry  23%  sulphur,  and  roughly 
69%  copper  and  iron.  Also  find  the  ratio  of  sulphur  to  base,  which 
here  is  3  to  1,  or  the  factor  3. 

Let  us  first  consider  the  sulphur.  Experience  shows  that  in 
regular  matte  smelting  we  can  depend  upon  a  loss  by  volatilization 
of  20  to  40%  sulphur.  We  take  25%  as  an  average,  and  thus  75% 
of  the  sulphur  is  left  to  form  matte.  This  is  75  Ib.,  and  multiplied 
by  the  factor  3  indicates  that  225  Ib.  Cu  and  Fe  together  are  needed 
to  satisfy  the  sulphur. 

The  slag  produced  is  calculated  by  dividing  the  weight  by  the 
per  cent,  expressed  decimally,  or  271  *ft  0.35  =  770  Ib.  Allowing 
0.5%  copper  for  the  slag  (and  in  good  work  it  should  not  exceed 
this)  the  weight  so  lost  is  4  Ib.,  leaving  96  Ib.  to  enter  the  matte. 
Subtracting  the  weight  of  this  copper  from  the  total  225  Ib.  of 
copper  and  iron  needed  for  the  matte,  we  get  129  Ib.  iron  entering 
the  matte,  out  of  the  total  302  Ib.  in  the  charge.  The  remainder 


318  THE    METALLURGY 

(173  lb.)  is  available  for  the  slag.  But  the  iron  existing  in  the 
charge  as  ferric  iron  is  reduced  to  ferrous  form,  and  we  have,  in 
the  ratio  of  atomic  weights,  56  parts  Fe  equal  72  of  FeO,  or  173  lb. 
Fe  equal  222  lb.  FeO.  To  this  we  add  the  158  lb.  CaO,  making  380 
lb.  of  the  two  bases.  Multiplying  the  silica  (271  lb.)  by  the  factor 
1.57  we  find  we  need  425  lb.  of  the  bases  FeO  and  CaO,  so  that 
we  have  a  deficit  of  45  lb.  Now,  since  the  limestone  consists 
approximately  half  of  CaO,  we  need  to  add  90  lb.  limestone  to 
the  charge,  making  in  all  390  lb.  as  the  required  amount.  Erase 
where  needed,  and  re-calculate  the  charge  throughout.  This  timv 
we  should  come  within  a  few  pounds  of  the  correct  amount.  As 
long  as  it  is  within  10  lb.  it  is  close  enough,  since  variations  in  the 
ores,  imperfect  weighing,  and  variation  in  the  amount  of  sulphur 
volatilized  easily  exceeds  such  differences.  When  we  have  learned 
the  actual  percentage  of  volatilization  by  experience  we  substitute 
it  for  that  above  assumed.  The  actual  percentage  of  copper  and 
iron  is  taken  in  the  same  way. 

The  grade  of  the  matte  in  copper  is  learned  from  the  ratio  of 
the  sulphur  (75  lb.)  to  the  copper  (96  lb.)  or  23  to  29</< .  In  the 
same  way  we  compute  from  the  respective  weights  the  percentage 
of  SiO2,  FeO,  and  CaO,  their  aggregate  being  90  per  cent. 

The  metallurgist  seldom  can  count  on  the  slag  and  matte  coming 
from  the  furnace  precisely  as  calculated.  There  is  a  little  variation 
due  to  the  causes  already  mentioned.  When  the  slag  from  a  newly 
calculated  charge  comes  down,  a  sample  should  be  taken  and  a 
determination  made  for  Cu,  SiO2,  FeO,  and  CaO.  It  should  be  pos- 
sible to  make  these  determinations  in  two  hours  and  to  correct  the 
charge  accordingly.  As  an  approximate  rule ;  to  increase  an  ingred- 
ient of  the  charge  a  given  percentage,  add  to  it  the  fractional  part 
expressed  by  its  ratio  to  the  remainder  of  the  100%.  Thus  if 
analysis  gives  33%  and  we  wish  to  increase  it  to  35%,  then  to  the 
2/33  of  271  =  16.4  lb.  add  33/67  or  i/2  of  16.4  lb.,  making  the  total 
silica  to  be  added  25  pounds. 

90.     PYRITE  MATTE  SMELTING. 

This  consists  in  treating  in  a  blast-furnace,  such  as  that  shown 
in  Fig.  131,  sulphide  ore  consisting  largely  of  pyrite  and 
chalcopyrite.  The  ore  carries  gold  and  silver,  which  are  recovered 
in  the  copper-bearing  matte  produced.  No  preliminary  roasting  is 
given  the  ore,  and  the  smelting  is  conducted  in  such  a  way  that 
70  to  80%  of  the  sulphur  is  burned  in  the  furnace  while  the 
remainder,  uniting  with  iron  and  the  copper,  forms  the  matte  which 


OF    THE    COMMON    METALS.  319 

acts  as  collector  for  the  gold  and  silver.  A  slag  is  formed  from 
the  silica  of  the  gangue  and  the  bases  of  the  ore  and  flux.  At  times 
when  the  quantity  of  base,  especially  iron,  is  large  it  is  necessary, 
in  order  to  make  a  suitable  slag,  to  add  silicious  ore.  The  matte 
and  slag  flowing  together  from  the  furnace  separate  in  the  fore- 
hearth. 

It  will  be  noticed  that  the  slow  and  expensive  preliminary 
roasting  of  the  sulphide  ores  is  obviated,  and  that  the  amount  of 
fuel  needed  is  small  (I1/-?  to  6%)  because  of  the  heat  developed 
by  the  burning  of  the  sulphide.  Pyrite  or  chalcopyrite  contains 
iron  that  is  available  both  for  matte  and  slag,  and  when  the  matte 
can  spare  it  for  the  slag  the  iron  serves  to  flux  the  silica  of  iron- 
free  ores  on  the  charge.  Iron  ore  or  limestone  acts  in  the  same 
way,  and  either  of  them,  though  generally  the  latter,  may  be  added 
for  the  purpose. 

An  iron  matte  alone  does  not  entirely  collect  the  gold  and  silver 
from  the  ore-charge,  and  it  has  been  found  that  copper,  to  the 
extent  of  0.5%  or  more,  should  be  present  to  insure  the  collection 
of  these  metals  in  the  matte.  The  slag  then  will  be  nearly  free 
from  the  precious  metals.  Copper,  therefore,  acts  as  an  efficient 
collector. 

In  result  of  burning  70  to  80%  of  the  sulphur  of  the  charge, 
there  remains  only  30  to  20%  to  form  matte.  The  remaining 
sulphur  first  takes  up  copper,  for  which  it  has  a  greater  affinity  than 
for  iron.  It  is  the  burning  off  of  the  large  amount  of  sulphur  that 
enables  one  to  dispense  with  roasting  and  to  diminish  the  amount 
of  matte  produced.  The  matte  produced  per  ton  of  ore,  or  the 
matte-fall,  may  be  expressed  as  a  percentage,  or  as  a  concentration 
of  so  many  tons  into  one  of  matte.  Thus,  with  a  production  of  200 
Ib.  matte  per  ton  of  ore,  we  have  a  10%  matte-fall,  or  a  concentra- 
tion of  10  into  1.  It  is  desirable  to  concentrate  the  ore  into  a  small 
bulk  of  matte.  To  show  how  such  concentration  is  effected,  both  in 
regular  matte  smelting  and  in  pyrite  smelting,  we  enter  upon  the 
following  considerations : 

In  regular  smelting  (with  a  charge  containing  8%  sulphur,  the 
volatilization-loss  being  25%,  and  the  matte  to  contain  25% 
sulphur),  we  have  from  100  Ib.  ore  75%  of  8%  =  6  Ib.  sulphur  to 
form  matte.  This  makes  24  Ib.  matte  and  results  in  a  concentration 
of  4.2  into  1. 

In  pyrite  smelting  with  a  charge  containing  30%  sulphur,  the 
volatilization  loss  being  80%  and  the  matte  still  to  contain  25% 
sulphur,  we  have  from  100  Ib  of  ore  20%  of  30%  =  6  Ib.  of  sulphur 


320  THE    METALLURGY 

to  form  matte.     This  makes  24  Ib.  of  matte-,  the  same  concentration 
as  in  the  regular  matte  smelting  just  specified. 

It  will  be  noted  that  the  percentage  of  volatilization,  or  the 
amount  of  sulphur  burned,  varies  with  the  charge.  It  is  low  when 
ily  roasted  or  oxidized  ores  are  used,  and  high  for  raw  or 
Inherent  in  pyrite  smelting  is  the  difficulty  of  regulating  this  loss. 
When  the  furnace,  which  is  burning  the  right  amount  of  sulphur, 
begins  to  run  slow  from  any  cause,  the  volatilization  may  increase 
to  the  point  of  burning  the  entire  content  of  sulphur,  so  that  no 
matte  is  produced.  On  the  other  hand,  when  the  furnace  begins 
to  run  fast,  much  matte,  low  in  copper  is  produced.  The  principal 
difficulty  in  pyrite  smelting  is  the  regulation  of  the  matte-fall. 

91.     REACTIONS  IN  PYRITE  SMELTING. 

Let  us  take  the  case  of  smelting  with  a  12-ft.  smelting  column. 
The  charge  is  composed  largely  of  pyrite,  with  chalcopyrite  and  a 
gangue  of  silica.  To  this  limestone,  for  flux,  has  been  added,  and 
a  small  percentage  of  coke,  which,  however,  does  not  interfere  with 
the  reactions. 

At  the  temperature  of  the  surface  of  the  charge  (250°C.)  the 
FeS2  reacts,  and  a  part  of  the  sulphur  escapes.  This  encountering 
air  entering  the  feed-door,  burns  with  the  characteristic  blue  flame 
j;o  S02,  while  Fe3S4  remains.  As  the  temperature  becomes  higher, 
jmroasted  ones,  ^especially  those  containing  pyrite.  3  defect 
5  ft.  down  from  the  top,  sulphur  is  expelled,  leaving  Fe!3. 

Seven  feet  down,  dissociation  continuing,  we  have  Fe3S4  =  4FeS 
+  Fe  or  a  condition  in  which  FeS  holds  Fe  in  solution,  so  that,  if 
a  sample  of  the  substance  in  a  molten  condition  were  withdrawn 
from  the  furnace,  we  would  find  Fe  separating  from  the  FeS  on 
cooling.  The  pyrite  at  this  zone  (925  to  950°C.)  begins  to  melt, 
and  the  descending ,  drops  meet  the  rising  air  of  the  blast,  and  the 
principal  reaction  of  the  furnace  takes  place  as  follows : 

4FeS     +     Fe     +•  130  =  5FeO     +     4SO2 
4X22,800  5X66,400     4X71,000=524,800 

This  is  equivalent  to  1874  pound-calories  per  pound  of  iron  present. 

Generally  not  all  the  FeS  is  oxidized,  a  part  escaping  the  action 
of  the  blast  and  entering  the  crucible.  Where  for  any  reason  the 
furnace  slows  down,  as  for  example  after  the  addition  of  a  large 
amount  of  silicious  material,  the  air  has  time  to  burn  the  FeS,  and 
little  or  no  matte  is  formed.  In  pyrite  smelting  the  degree  of 
concentration  depends  upon  the  amount  of  air  blown  into  the 
furnace,  and  the  more  silica  we  have  in  the  slag,  the  greater  will 


OF    THE    COMMON    METALS.  321 

be  the  concentration.  In  a  well-working  furnace  300  Ib.  air  is 
needed  to  1  Ib.  sulphur,  and  the  silica  would  approximate  32%. 
If  the  concentration  is  low,  we  may  add  silica  or  quartz,  and  if 
high,  discontinue  feeding  it.  Considered  in  another  way,  the  pyrite 
furnace  chooses  its  own  slag. 

Below  the  preparatory  zone,  FeO  and  CaO  (if  present)  are 
transformed  into  slag,  uniting  with  the  silica.  The  unmelted 
material  keeps  the  charge  open  and  porous,  and  the  ore  column, 
instead  of  resting  on  a  bed  of  burning  coke  in  the  crucible,  as  when 
coke  is  used,  rests  upon  a  net-work  of  quartz  pieces  that  have  thus 
far  escaped  being  slagged.  The  net-work  of  unfused  pieces  extends 
from  the  crucible  upward  to  the  zone  of  fusion  of  the  iron  sulphide, 
7  ft.  above  the  tuyeres.  At  the  sides  the  net-work  is  moving  slowly 
downward  and  sustaining  the  portion  that  is  descending  regularly. 
At  5  to  6  ft.  above  the  tuyeres  is  found  the  CO,,  the  result  of  the 
combustion  of  the  coke  near  the  tuyere-zone. 

We  may  consider  that  the  chalcopyrite  loses  some  sulphur,  but 
is  little  changed,  because  of  the  superior  affinity  of  copper  for 
sulphur  as  compared  with  iron,  and  in  consequence  it  melts,  forms 
matte,  falls  in  drops  to  the  crucible,  and  at  the  same  time  seeks  and 
collects  the  gold  and  silver  of  the  charge. 

92.     PYRITE  SMELTING  IN  TWO  STAGES. 

For  low-grade  ores,  carrying  2%  copper  for  example,  a 
concentration  of  ten  into  one  gives  matte  of  20%  copper.  By 
roasting  the  matte,  or  by  smelting  it  pyritically,  it  is  possible  to 
increase  the  grade  to  40%  or  more,  and  this  product  can  be  treated 
in  the  copper-converter  and  brought  to  the  grade  of  blister  copper. 
An  ore  containing  5%  copper  can  be  smelted  to  give  a  matte  of  40% 
copper,  so  that  the  second  smelting  with  the  additional  expense 
can  be  omitted.  Now  while  it  would  not  pay  to  smelt  a  copper  ore 
of  a  grade  as  low  as  2%  copper,  for  the  copper  alone,  if  the  ore 
contained  gold  and  silver  the  recovery  of  these  metals  would  justify 
the  expense  of  smelting. 

Two-stage  smelting  at  Ducktown,  Tennessee. — At  the  works  of 
the  Tennessee  Copper  Co.,  the  process  is  a  two-stage  one,  the  ore 
being  smelted  to  give  matte  of  10%  Cu  ('ore-smelting'),  this  matte 
being  re-treated  in  another  furnace  to  produce  a  35  to  45%  matte 
('matte-smelting').  The  slag  from  the  second  furnace,  not  yet 
sufficiently  clean  to  throw  away,  is  re-melted  in  the  first  furnace. 
The  second  matte  is  then  converted  to  blister  copper  of  88  per  cent. 

The  furnaces  (of  the  type  shown  in  Fig.  131),  are  180  in.  long 


322 


THE    METALLURGY 


by  56  in.  wide  at  the  tuyere  level,  and  have  large  circular  fore- 
hearths,  16  ft.  outside  diameter  by  5  ft.  deep,  lined  with  chromite 
brick  to  resist  the  corrosive  action  of  the  low-grade  matte  produced. 
The  first,  or  ore-furnace,  treats  400  tons  of  charge  daily,  with  a 
coke-consumption  of  2.1%.  The  second  or  concentration  furnace 
smelts  280  to  300  tons  of  matte,  using  3.5%  coke.  The  following 
charge-sheet  gives  details  regarding  the  charge  and  the  matte 
produced  in  both  the  stages,  and  shows  how  such  charges  are 
computed  in  pyrite  smelting. 

93.     CALCULATION  OF  CHARGE  IN  PYRITE  SMELTING. 

In  these  calculations  the  quantity  of  coke  is  so  small  that  no 
computation  is  required  for  the  ash.  The  quantity  of  base  in  the 
ore  is  so  large,  and  the  silica  is  so  low  that  it  has  been  necessary 
to  add  silicious  material  (in  this  case  quartz-rock)  to  the  charge 
in  order  to  obtain  a  slag  of  35%  silica.  The  problem  is  to  compute 
the  amount  of  quartz  to  be  added  to  give  such  a  slag. 

CHARGE    SHEET   I.      ORE-SMELTING   FURNACE. 


Name  of  Ore 

Weight 

H20 

Cu 

SiC-2 

Fe  4-  Mn 

CaO+MgO 

s 

Wet 

Dry 

% 

lb. 

% 

lb. 

% 

lb. 

% 

lb.    I   %   ;  lb. 

Polk  Co  

1000 

2.4 

24 

20.7 

207 

34.2 

342 

9.2 

92 

30,4 

9,04 

Burra-Burra  

3000 

2.1 

63 

9.4 

282 

38.0 

1140 

8.3 

249 

30.3 

909 

Quartz  

700 

97.0 

679 

Coke  

100 

Q7 

1  1fiS 

1400 

911 

1113 

Cu  in  Slag 

6 

Volatilized=890 

"   "  Matte                81 

In  Slag       =  43  _933 

Matte   =         180 

Matte. 

Fe  +  Cu  =  65% 
S  =  25% 
Cu  +  Fe 


Slag. 

Si02   =  35.0% 

FeO  +  CaO    =  55.0% 

S   =    1.3% 

Cu   —    0.2% 

Slag-  —  1168  -=-  0.35  —  3300 

qc 

Factor   ~  =  0.636 
oo 

We  represent  on  the  charge-sheet  the  ores  that  are  to  be  run, 
using  amounts  in  accord  with  the  rate  at  which  the  respective  ores 
are  supplied  (1000  lb.  Polk  county  ore  and  3000  lb.  Burra-Burra 
ore).  Experience  shows  that  for  such  a  charge  and  for  the  quantity 
of  ferrous  iron  and  lime  present,  we  may  enter  the  quantity  of 
silicious  material  as  700  lb.  We  use  a  slag  of  35%  SiO2  and  55%"" 
FeO  and  CaO,  making  in  all  90%.  With  this  slag,  experience  shows 


OF   THE    COMMON   METALS.  323 

we  may  figure  on  1.3%  sulphur  and  0.2%  copper  for  this  low-grade 
matte.  A  little  zinc,  when  that  element  is  present  in  the  charge,  also 
enters  the  slag.  This  is  shown  to  be  0.3  per  cent. 

The  matte  is  assumed  to  contain  25%  sulphur,  65%  copper  and 
iron,  and  1.7%  zinc.  We  estimate  that  80%  of  the  zinc  will  be 
volatilized.  It  is  understood  that  in  smelting  other  ores  than  these, 
the  actual  quantities  of  the  different  elements  in  the  matte  and  slag 
will  be  determined  and  those  figures  substituted  for  the  ones  above. 

The  percentage  of  ingredients  of  the  charge  is  written  and 
carried  out  in  the  respective  columns,  and  the  columns  are  added. 

Beginning  with  the  sulphur,  of  the  1113  Ib.  present,  we  have : 

Pounds. 

Sulphur  volatilized  (80%  of  the  total)  =  890 
in  the  slag  (1.3%  of  3300  Ib.)  =  43 
left  for  the  matte  =  180 


1113 

Of  the  copper,  0.2%  of  3300  Ib.  or  6  Ib.  goes  into  the  slag,  leaving 
81  Ib.  for  the  matte.  Multiplying  the  sulphur  for  the  matte,  180  Ib., 
by  the  factor  2.6  we  get  the  total  Fe  and  Cu  needed  for  matte, 
468  Ib. ;  substracting  Cu  for  matte,  81  Ib. ;  leaves  Fe  for  matte,  387 
pounds. 

But  the  total  iron  in  the  charge  is  1482  Ib.,  so  that  we  have : 

Pounds. 

Fe  in  matte 387 

Fe  left  for  slag ,   1095 


Total 1482 

The  iron  in  the  slag  occurs  as  FeO,  so  that  we  must  take  1408  Ib. 
FeO,  equivalent  to  the  1095  Ib.  Fe.  Adding  to  this  the  CaO,  341  Ib., 
we  get  FeO  +  CaO  =  1749  Ib.,  which  multiplied  by  the  factor  0.636, 
gives  Si02  =4112  Ib. 

Pounds. 

Actual  silica  in  charge 1168 

Silica  needed.  .  ,   1112 


Silica  in  excess 56 

By  erasing  the  trial  amount  700  Ib.  of  quartz,  and  substituting 
650  Ib.,  then  re-calculating  the  charge,  we  get  an  approximation 
within  10  to  20  Ib.,  which  is  accurate  enough  for  practical  purposes. 


324 


THE    METALLURGY 


The  percentage  of  copper  in  the  matte  is  computed  according  to  the 
proportions  195  :81 :  :25%  :10.4  per  cent. 

94.     CALCULATION  OF  MATTE  CHARGE. 

The  matte  charge  is  run  with  slag  from  the  converting  operation, 
and  quartz  ore  yi  sufficient  quantity  to  produce  a  slag  of  the  same 
composition  as  that  of  the  ore-charge,  except  that  it  has  1%  sulphur 
and  0.7%  copper. 

CHARGE-SHEET    II.       MATTE-CONCENTRATING    FURNACE. 


Name  of  Ore 

Weights 

H20 

Cu 

S1O2 

Fe  +  Mn 

CaO+  Mno 

S 

Wet 

Dry 

% 

Ib. 

% 

Ib. 

% 

Ib. 

% 

Ib. 

% 

Ib. 

Matte             

3000 

10.0 

300 

55.0 

1650 

?50 

750 

Quartz  

1400 

97.0 

1358 

Limestone  

200 

4.0 

8 

53.0 

106 

Convert.  Slag  

400 

2.0 

8 

30.0 

120 

55.0 

220 

1.0 

4 

1.0 

4 

Coke  

100 

308 

1486 

1870 

110 

754 

Cu  in  Slag     =         30 

Volatilized  530 

"    "  Matte  =      278 

In  Slag           43    573 

181 

Matte. 
Fe  +  Cu  =  65% 

S  =  25% 

CU  +  Fe  =  2.6 


Slag. 

Si02   =  35.0% 

FeO  +  CaO   =  55.0% 

S   =     1.0% 

Cu   =    0.7% 

Slag  =  1486  -r-  0.35  =  4300  Ib. 

Factor  |i  =  0.636 
bo 

The  charge  is  estimated  as  in  'Charge-sheet  I',  with  a 
volatilization  of  70%  of  the  sulphur,  as  experience  has  shown  the 
result  commonly  to  be.  In  the  matte  we  have: 

Pounds. 

S    181  =  25.0 

Cu     278  =  38.4 

(181  X  2.6)-278  =  Fe 193  =  26.6 

652  =  90.0 

Proceeding  with  the  calculation  for  the  quartz  we  have : 

Pounds. 

Total  iron 1870 

Iron  in  matte.  .  193 


Iron  for  slag 

or  FeO  =  2156,  and  the  total  base  is  2266  Ib. 


1677 


OF    THE    COMMON    METALS.  325 

This  gives  the  silica  needed,  2266X0.636  =  1509  Ib.  But  we 
have  already  1486  Ib.,  and  the  difference  may  be  made  up  by 
increasing  the  quartz  20  pounds. 

95.     MATTE  CONCENTRATION. 

In  an  example  above,  we  obtained  a  matte  10%  in  copper,  and 
with  copper  low  in  the  charge,  the  percentage  may  be  even  less. 
The  matte  is  of  too  low  grade  to  ship  away,  or  to  bring  to  the  grade 
of  blister-copper  in  a  converter.  It  must  be  concentrated  to  one 
of  higher  grade.  If  we  were  to  heap-roast  the  matte  and  then  smelt 
it  with  silicious  ore  in  a  blast-furnace,  we  should  obtain  a  small 
quantity  of  matte  of  a  high  grade. 

There  is,  however,  the  expense  and  delay  of  the  roasting  to 
consider,  and  it  has  been  sought  to  smelt  the  matte  raw,  with 
silicious  ore,  with  the  idea  of  burning  off  the  sulphur  in  the  blast- 
furnace. In  regular  matte-smelting,  were  this  attempted,  the  matte 
would  run,  though  little  diminished  in  quantity  and  little  changed 
in  grade,  but  by  the  new  method,  using  little  fuel,  an  abundant 
blast,  and  silicious  slag,  the  concentration  can  be  obtained. 

If  the  matte  is  to  be  transported  for  treatment  it  should  be  at 
least  50%  copper.  If  of  40%,  however,  it  can  be  treated  in  the 
converter  to  bring  it  up  to  the  grade  of  blister  copper.  Where 
it  is  difficult  even  with  two-stage  smelting  to  bring  the  matte  to 
this  concentration,  a  lower-grade  matte  may  be  made  and  shipped 
away  for  treatment  elsewhere. 

96.     SLAG  DISPOSAL. 

The  slag  from  a  blast-furnace,  being  a  waste  material,  is 
disposed  of  in  the  cheapest  way  possible.  In  the  case  of  small 
furnaces,  as  it  flows  from  the  fore-hearth,  it  is  caught  in  wheeled 
slag-pots  (slag  carts),  Fig.  135,  that  are  taken  to  the  edge  of  the 
slag-dump  when  filled  and  poured.  As  the  dump  grows  the  expense 
increases,  and  large  slag  cars.  Fig.  137,  are  used.  The  cars  are 
moved  either  by  horses,  or  an  industrial  locomotive,  or  by  trolley. 

Another  cheap  and  favorite  way  is  to  granulate  the  slag.  To 
do  this  a  cast-iron  launder  is  arranged  to  receive  the  slag  as  it 
falls  from  the  spout  of  the  fore-hearth.  The  launder  has  a  grade 
of  1  in.  to  the  foot,  and  through  it  water  is  made  to  flow  constantly. 
In  addition,  a  horizontal  flattened  jet  of  water  strikes  the  falling 
slag,  instantly  cooling  and  breaking  it  into  granules  of  various 


326 


THE    METALLURGY 


sizes  averaging  one-sixteenth  of  an  inch  in  diameter.     The  flow  of 
water  carries  the  slag  to  the  dump. 

97.     REVERBERATORY  MATTE-SMELTING. 

While  the  blast-furnace  in  general  is  the  cheapest  means  of 
smelting  copper-bearing  ores  in  coarse  or  lump  form,  one  objection 
to  it  is  that  the  strong  blast  necessarily  used  may  carry  away  5 


Fig.    137.      ELECTRIC'  TROLLEY   SYSTEM. 

to  10%  of  the  fine  dusty  ore.  Such  ore  may  be  settled  as  flue-dust, 
in  flues  and  dust-chambers,  and  made  into  briquettes  and  re-smelted, 
but  this  is  an  additional  expense  to  be  avoided  if  possible.  In  fact, 
ore  or  concentrate,  in  fine  condition,  is  better  treated  in  rever- 
beratory  furnaces.  When  putting  the  charge  into  the  furnace  the 
stack-damper  is  closed,  and  when  the  dust  from  the  dropping  ore 
had  subsided  it  is  again  opened.  The  ore  is  not  then  disturbed. 
If  raw  sulphide  were  melted  in  such  a  furnace,  no  sulphur  would 
be  expelled;  hence  before  such  treatment,  sulphide  ore  must  be 
roasted,  generally  in  one  of  the  types  of  mechanical  roasters.  The 
output  of  a  reverberatory  is  much  less  than  that  of  a  blast-furnace, 
and  a  larger  proportion  of  fuel  is  needed,  though  it  may  be  of  a 
cheaper  kind.  Thus  we  use  in  •  reverberatory  smelting,  30%  of 
bituminous  coal  as  against  10%  of  coke  in  regular  blast-furnace 
smelting,  or  as  little  as  1.5  to  6%  in  pyrite  smelting.  We  may 


OF   THE    COMMON   METALS.  327 

say,  as  far  as  fuel  is  concerned,  that  the  30%  fuel  at  $2.50  per  ton 
would  balance  the  10%  coke  at  $7.50  per  ton. 

98.     THE  REVERBERATORY  MATTING-FURNACE. 

Fig.  138  represents  an  elevation  and  Fig.  139  a  sectional  plan 
of  a  large  reverberatory  matting-furnace,  having  a  hearth  37.5  ft. 
long  by  15  ft.  wide  with  a  fire-box  7  by  8  ft.,  or  of  56  sq.  ft.  area. 
Over  the  furnace  are  six  hoppers  for  charging  the  ore,  and  one 
double  hopper  from  which  coal  is  charged  as  needed,  to  the  fire- 
box. The  front  of  the  furnace  is  the  flue-end;  the  back  is  the 
fire-box  end.  The  lager  portion  of  the  charge  is  fed  near  the 
latter,  the  hotter  part  of  the  furnace.  The  interior  line  of  the 
furnace  is  shown  by  dotted  lines  in  the  elevation  that  shows  also 
the  fire-box,  the  bridge  or  wall  separating  the  hearth  from  the 
fire-box,  the  hearth,  and  outlet-flue.  As  shown  in  the  plan,  the 
bridge  is  furnished  with  a  double  cast-iron  plate,  called  'conker 
plate.'  This  serves  to  support  and  strengthen  the  bridge.  There 
are  two  doors  at  either  side  of  the  furnace  by  which  access  is  had 
for  repairs,  but  these  commonly  are  kept  tightly  closed.  At  the  side 
near  the  middle  is  noticed  the  side  skimming-doors.  Slag  is  skimmed 
also  from  the  surface  of  the  molten  charge  at  the  front-end,  through 
a,  door  12  in.  above  the  hearth-bottom. 

At  the  back  of  the  furnace,  between  the  two  doors,  is  noticed 
the  matte  tap-hole.  It  is  set  low  to  drain  the  furnace  of  the  entire 
content  when  required.  By  means  of  it  the  matte  is  withdrawn 
after  the  supernatant  slag  has  been  skimmed  from  the  charge. 

At  the  height  of  the  hearth-bottom  are  the  grate-bars,  and  below 
them  the  ash-pit,  5  ft.  deep.  The  fire  is  'grated'  or  cleaned  twice 
in  24  hours.  The  doors  of  the  ash-pit  are  opened  and  men  enter  to 
remove  the  ashes  and  clinkers  from  the  grate,  using  a  long,  stout, 
steel  bar  for  the  purpose.  The  material  falls  into  a  low  car  on  a 
track  shown  in  plan,  Fig.  139.  The  fire  having  been  grated,  the 
car  is  removed,  the  ash-pit  doors  are  closed  and  tightly  luted  with 
clay,  and  air  from  a  fan-blower  is  admitted  from  a  pipe  (shown  in 
the  plan)  to  the  ash-pit.  This  gives  an  under-grate  pressure,  burns 
the  coal  rapidly,  and  decreases  the  necessary  stack-draft  and  the 
consequent  sucking  of  cold  air  into  the  furnace.  The  stack  is 
separate  from  the  furnace,  the  smoke  entering  by  a  sloping  flue. 

^Operation  of  the  furnace.— While  a  blast-furnace  is  fed  with 
successive  charges,  the  reiverberatory  receives  a  charge  of  many 
tons  at  one  time.  The  roasted  ore  for  the  reverbatory  is  stored 
in  the  hoppers  above  the  furnace,  and  drops  into  the  furnace  when 


THE    METALLURGY 


OF    THE    COMMON    METALS. 


329 


the  charge-opening  is  uncovered  and  the  hopper-slide  withdrawn. 
If  the  ore  were  to  drop  directly  upon  the  hearth  the  cold  charge 


Fig.   139.     REVERBERATORY  SMELTING  FURNACE    (PLAN). 

77, 


330  THE    METALLURGY 

would  adhere  to  it  and  would  be  slow  to  heat  and  melt.  It  is 
customary  to  retain  a  pool  of  matte  in  the  furnace  upon  which  the 
ore  drops  and  floats  out.  At  the  time  of  charging,  the  fire  is 
cleaned  or  grated.  Firing  now  proceeds  vigorously  until  the  charge 
is  melted.  This  work  takes  several  hours. 

The  sulphur  of  the  charge  unites  with  the  copper  and  iron  until 
its  needs  are  satisfied.  The  matte  thus  formed,  separating  in  drops, 
absorbs  the  precious  metals  contained  in  the  ore,  and  by  the  greater 
specific  gravity  than  that  of  the  slag,  penetrates  downward  to  the 
hearth.  The  silica  of  the  gangue  unites  with  base,  such  as  Fe07 
CaO,  and  A1203,  and  forms  a  fusible  slag  that  floats  as  a  separate 
layer  upon  the  matte. 

When  ore  is  roasted  a  part  of  the  iron  is  oxidized  to  the  ferric 
state,  and  when  the  charge  is  fused  the  ferric  iron  acts  upon  the 
unroasted  ferrous  sulphide  according  to  the  following  equation : 

FeS  +  3Fe203  +  7Si02  =  7FeSiO8  +  S02 

When  ferric  iron  is  present  in  the  ore,  sulphur  is  eliminated 
often  to  the  extent  of  25  to  33%,  and  we  find  less  matte  than 
would  be  present  if  this  reaction  did  not  take  place.  The  iron, 
thus  reduced  to  ferrous  form,  enters  the  slag. 

The  skimming-doors,  one  or  both,  are  opened,  and  the  slag 
is  skimmed  off  by  means  of  long-handled  rabbles.  As  the  slag  is 
removed  it  falls  into  slag-carts  that  are  wheeled  away  to  the  dump. 
The  side-doors  of  the  furnace  are  now  opened,  and  the  interior 
walls  repaired,  especially  at  the  bridge  end  of  the  furnace  where 
they  have  been  eaten  out  by  the  action  of  the  molten  slag.  The 
repair-work  or  fettling  is  done  by  skilfully  throwing  sand  from 
the  side-door  across  the  furnace  so  that  it  banks  up  against  the 
opposite  wall.  The  repair  is  also  made  by  placing  the  sand  upon 
the  desired  spot  by  means  of  a  long-handled  paddle  or  spoon.  Balls 
of  ganister  are  brought  to  the  spot  and  pressed  against  the  side 
by  the  same  tool.  The  side-doors  are  now  closed  and  luted  tightly 
with  a  clay  mortar,  and  the  charge,  already  in  the  hopper,  is  put 
into  the  furnace.  The  operation  cools  the  furnace  as  does  also  the 
grating  of  the  fires. 

The  matte,  when  tapped,  runs  into  sand-molds  or  depressions 
made  in  the  sand-floor  of  the  furnace-house.  Another  way  of 
handling  is  to  draw  the  matte  into  ladles  that  are  operated  by 
an  overhead  traveling  crane,  by  means  of  which  it  is  transferred 
to  the  converter. 


OF    THE    COMMON    METALS.  331 

99.  REVERBERATORY  SMELTING  (WELSH  PROCESS). 

The  process  consists  in  treating  copper  ore  (sulphide  and  oxide 
as  well  as  silicious  ore)  by  a  series  of  roastings  and  fusions  to  raise 
the  grade  of  the  copper  finally  to  blister  copper,  which  subsequently 
is  refined.  It  possesses  the  advantage  that  a  variety  of  ores  both 
coarse  and  fine  can  be  treated  in  a  few  reverberatory  furnaces 
without  large  investment  in  plant.  For  a  small  tonnage,  it  is 
possible  to  produce  on  the  spot  metallic  copper. 

We  may  divide  the  process  into  five  operations: 

(1)  'Calcining'  the   ore:     Sulphide   ore   containing   5   to   15% 
copper  is  roasted  in  a  hand-reverbatory  roaster   (See  Fig.  31  and 
32)  until  not  more  than  5%  sulphur  is  left. 

(2)  Fusion  of  ore :    The  roasted  ore  is  charged  in  a  reverbatory 
furnace  with  such  oxidized  copper  ore  as  is  available,  and  melted. 
The   sulphur   contained   in   the   roasted   ore,   with   the   copper   and 
some  of  the  iron,  forms  a  matte  of  35%  copper,  called  'coarse  metal.' 
The  silica,  uniting  with  the  ferrous-oxide  not  taken  by  the  matte, 
and   also   with   the    alumina,    and   the   alkaline-earth   bases,    forms 
a  slag,  fusible  at  the  high  temperature  of  the  furnace.     The  molten 
bath  boils  from  the  escape   of  sulphur  oxides  resulting  from  the 
reaction  of  the  ferric  iron  upon  unroasted  ferrous  sulphide,  or  from 
the  decomposition  of  barium,  lead,  or  zinc  sulphide  by  silica,  thus : 

BaSO4  +  SiO2  =  BaSiO3  +  S03  N  dfy 

PbSO4  +  SiO2  =  PbSiO3  +  S03 

ZnS04  +  SiO2  =  ZnSiO8  +  S03 

The  sulphuric  anhydride  escapes  as  a  gas,  and  upon  meeting 
the  moisture  of  the  air  at  the  top  of  the  stack,  changes  to  a  white 
fume  of  H2S04. 

(3)  Calcining  coarse-metal:     The  coarse  metal  or  matte  that 
is  run  from  the  reverberatory  furnace  in  operation   (2)   into  sand 
beds,  is  crushed  to  pass  a  5-mesh  screen  and  fed  to  another  hand- 
roaster.     Rich  sulphide  of  20  to  70%   copper  also  is  crushed  and 
added  to  the  charge.     The  whole  is  roasted  until  it  contains  not 
more  than  5%  sulphur. 

(4)  Second  reverberatory  fusion :     The  roasted  material,  now 
of  35  to  50%  copper,  is  charged  into  a  fusion-furnace  with  oxidized 
ores   containing  20   to   70%    copper.     When  the   charge   is  melted 
there  results  a  matte  of  75%  copper,  called  'white  metal,'  composed 
chiefly  of  copper  sulphide.     As  before,  the  silica  contained  in  the 
ore  added  to  the  charge  unites  with  the  ferrous  iron  and  other  bases 
to  form  slag.     This  slag,  however,  having  been  made  from  such 


THE    METALLURGY 

rich  material,  contains  much  copper  and  is  not  to  be  thrown  away 
but  returned  to  another  charge  in  the  fusion  furnace  of  opera- 
tion (2). 

(5)  'Roasting'  and  formation  of  blister  copper:  The  white 
metal  is  charged  in  large  pieces,  as  broken  when  removing  from  the 
sand  molds,  into  a  reverberatory  fusion-furnace  where  it  is  piled 
in  an  open  fashion  particularly  near  the  bridge.  It  is  fired 
gradually  several  hours  with  an  oxidizing  flame.  A  small  supply 
of  air  is  admitted  at  a  number  of  ports  ,or  openings  2.5  in.  sq. 
in  the  roof  over  the  fire-bridge,  and  at  the  sides  of  the  furnace 
near  the  bridge.  The  operation  is  called  'roasting.'  The  oxidizing 
flame,  acting  at  the  surface  of  the  lumps  and  upon  the  drops 
trickling  down,  converts  a  portion  into  cuprous  oxide,  so  that  we 
have  present  copper  both  as  sulphide  and  as  oxide.  Finally  the 
heat  is  raised  and  the  whole  charge  is  melted  down.  Now  occurs 
the  endothermic  reaction  :  ^ 

2Cu20       +       Cu2S      ===       6CuS02 
2X42,000  20,200  71,6bo=  -33,200 

Copious,  fumes  of  S02  issue  from  the  boiling  surface  of  the  molten 
charge.  Slag  rich  in  copper  is  produced,  getting  the  silica  partly 
from  the  interior  walls  of  the  furnace,  partly  from  silicious  but 
entirely  oxidized  ore  that  has  been  added  to  supply  silica.  The 
slag  is  returned  to  operation  (4).  Finally  the  blister-copper  is 
tapped.  The  metal  obtains  this  name  from  the  fact  that,  upon 
cooling,  occluded  gas  seeking  to  escape  from  the  molten  metal, 
forms  blisters  on  the  surface  of  the  pigs  of  metal. 

The  blister-copper  now  contains  98%  copper,  but  also  impurities 
that  must  be  removed  to  make  it  suitable  for  market.  The  refining 
process  is  described  elsewhere.  In  case  the  copper  contains  gold 
and  silver,  taken  from  the  ores  that  supplied  the  copper,  it  is 
customary  to  re-melt  it,  pole  it  to  remove  copper  oxide,  and  to 
cast  it  into  anodes  for  electrolytic  refining. 

100.     BEST-SELECTED   COPPER. 

Where  copper  ores  are  impure,  and  at  the  same  time  where  it 
is  desired  to  obtain  a  superior  grade  of  refined  copper,  the  Welsh 
process  is  modified  as  follows : 

A  portion  of  the  white-metal,  produced  in  operation  (4)  is 
ground  to  5-mesh  and  roasted  or  calcined  so  that  part  of  the  Cu2S 
is  oxidized  to  Cu20..  This  roasted  matte  is  put  into  a  fusion-furnace 
and  melted  down  quickly  with  a  large  proportion  of  white-metal 
in  lump  form.  A  reaction  occurs  between  the  Cu20  present  and 


OF    THE    COMMON    METALS.  333 

some  of  the  Cu2S,  and  a  small  part  of  the  charge  (one-fifteenth  of 
the  whole)  is  reduced  to  metal.  The  metal  takes  the  impurities 
of  the  charge  such  as  arsenic,  antimony,  and  tellurium,  and  also 
the  gold  if  present,  while  the  unreduced  matte,  freed  from  these 
impurities,  is  pure.  After  skimming  the  charge  the  content  of  the 
furnace  is  tapped  into  sand-molds.  In  the  first  few  molds  is  found 
a  layer  of  copper  that  has  been  produced.  It  lies  beneath  the  matte 
that  forms  most  of  the  pig  and  is  called  'bottoms.'  The  pure  matte 
thus  obtained  is  treated  as  in  process  (5),  and  a  blister-copper, 
comparatively  free  from  impurities,  is  obtained. 

Treatment  of  the  bottoms. — The  impure  bottoms  are  re-melted 
and  cast  into  anodes  for  electrolytic  refining,  in  which  the  impurities 
and  gold  are  separated  from  the  copper ;  or  they  may  be  formed 
into  an  inferior  grade  of  copper  (casting  copper)  as  follows:  A 
charge  is  put  into  the  blister-furnace  as  in  operation  (5),  consisting 
of  14,000  Ib.  of  75%  roasted  white-metal,  21,000  Ib.  raw  white-metal, 
8000  Ib.  bottoms,  and  1000  Ib.  silicious  ore.  This  is  melted  down, 
and  then  is  added  6000  Ib.  more  roasted  white-metal.  The  charge 
aggregates  50,000  Ib.  This  is  treated  precisely  like  the  regular 
blister  charge,  but  it  yields  a  higher  percentage  of  copper. 

101.     THE  DIRECT  PROCESS. 

This  is  a  modification  from  the  Welsh  method  of  producing 
blister  copper.  Instead  of  'roasting'  the  matte  or  white-metal  in 
lump  form  in  the  blister-furnace  or  process  (5),  a  portion  is  ground 
to  5-mesh  size  and  calcined  or  roasted  in  a  separate  roasting-furnace, 
of  either  the  hand  or  the  mechanical  type. 

Into  a  melting  furnace,  called  the  'blister-furnace,'  is  charged 
14,000  Ib.  of  the  roasted  white  metal,  still  retaining  4  to  6%  sul- 
phur, 3500  Ib.  raw  unroasted  white  metal,  4000  to  8000  Ib.  slag 
from  a  former  charge,  and  600  Ib.  silicious  ore  to  unite  with  the 
FeO  and  other  base  present.  When  this  has  been  melted,  6000  Ib. 
more  roasted  matte  is  added,  making  a  total  of  23,500  Ib.  matte 
charged.  When  all  is  fused,  the  reaction  begins.  The  surface 
of  the  charge  is  seen  to  be  seething  and  boiling,  and  escaping  bubbles 
of  gas  are  set  free  according  to  the  following  reaction : 

2Cu20  +  Cu2S  =  6Cu  +  S02 

The  bath  is  then  skimmed  to  remove  the  slag,  which  contains  copper 
oxide,  reserved  in  part  for  the  next  charge  and  in  part  sent  back 
to  stage  (4)  of  the  Welsh  process.  From  the  charge  here  specified 
there  is  produced  75  pigs  weighing  230  Ib.  each,  or  17,250  Ib. 
blister  copper,  and,  also,  23  pots  of  slag  weighing  400 -lb.  or  9200 
Ib.  containing  12  to  15%  copper. 


334  THE   METALLURGY 

102.     REVERBERTORY    MATTE-SMELTING    AT    ANACONDA, 

MONTANA. 

There  are  eight  reverberatory  melting-furnaces  at  this  plant, 
that  are  intended  for  the  production  of  a  matte  of  40%  copper 
from  roasted  concentrate.  The  matte  is  treated  in  converters. 
Pig.  140  and  141  represent,  in  elevation  and  plan  respectively,  one 
of  the  reverberatory  melting-furnaces  at  the  Washoe  plant  of  the 
Anaconda  Copper  Mining  Co.  Two  boilers  are  heated  by  the  waste 
gas  from  the  furnace,  and  develop,  together,  600  hp.  The  furnace 
treats,  on  an  average,  275  tons  in  24  hours,  producing  a  4:0%  matte 
with  a  concentration  of  4  into  1. 

The  charge  consists  of  hot  roasted  ore  ('calcines')  from  the 
McDougall  roasters,  Fig.  43,  and  by  analysis  is  shown  to  be 
composed  as  follows:  Cu  9%,  FeO  24.4%,  CaO  2.9%,  S  8%,  Si02 
26%.  Every  80  minutes  a  charge  of  15  tons  is  dropped  into  the 
furnace  from  the  two  hoppers  (shown  by  full  lines  in  the  elevation) 
near  the  fire-bridge.  This  falls  upon  the  bath  of  molten  matte  and 
slag  that  the  furnace  contains.  It  spreads  in  all  directions,  and 
much  of  it  floats  gradually  toward  the  front.  It  readily  melts  by 
contact  with  the  molten  slag  and  matte  below  and  the  flame  above. 
In  this  great  reservoir  of  heat  there  is  but  little  variation  in 
temperature,  and  the  flame  is  transparent. 

The  fuel  is  soft  coal,  21%  of  the  charge  or  56  tons  daily,  upon 
110.7  sq.  ft.  of  grate  surface,  or  40  Ib.  per  sq.  ft.  per  hour.  Every 
four  hours  45  to  50  tons  of  slag  is  removed  in  15  minutes  from 
the  furnace  and  allowed  to  flow  from  the  front  door  in  a  thick 
stream.  It  is  granulated  by  a  strong  horizontal  stream  of  water  as 
it  falls  into  the  waste-launder.  The  water  sweeps  it  away  to  the 
dump  several  hundred  yards  from  the  furnace.  The  matte  is  kept 
at  a  nearly  uniform  level,  ten  tons  being  tapped  out  at  a  time, 
while  the  total  amount  in  the  furnace  is  100  to  150  tons. 

The  action  of  the  slag  upon  the  furnace  is  to  erode  or  scour 
it,  but  because  of  the  width,  the  sides  are  less  acted  upon  than 
would  be  the  case  in  a  narrow  furnace.  To  repair  the  furnace,  as 
is  done  monthly,  both  slag  and  matte  are  drawn  off  completely,  then 
the  side-doors  are  opened,  and  sand  is  thrown  against  the  sides 
where  eaten  away  by  the  slag.  They  thus  are  protected  against  the 
inroads  of  the  slag.  A  furnace  runs  six  or  eight  months  and  then 
has  to  be  shut  down  for  the  thorough  repair  of  the  roof,  walls,  and 
bridge.  These  parts  are  of  silica  brick,  the  walls  being  30  in.  thick, 
the  roof  15  in.  Silica  bricks  are  practically  infusible  but  expand 


OF    THE    COMMON    METALS. 


335 


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336  THE    METALLURGY 

on  heating,  so  that  allowance  is  made  by  leaving  transverse  slits 
in  the  roof.  These  close  when  the  furnace  is  at  full  heat.  An 
important  point  in  efficient  working  is  to  have  the  outlet-flue,  or 
'neck/  of  the  proper  size.  It  must  be  large  to  insure  good  draft, 
and  yet  retain  the  flame  in  the  furnace.  In  the  furnace  shown  it 
is  60  by  38  in.  or  16  sq.  ft.  area.  The  ash-pit  slopes  steeply  at 
the  back  to  the  ash-launder  which  carries  a  constant  liberal  stream 
of  water.  The  ash  and  cinder  fall  into  the  launder  and  are  at  once 
carried  away.  Hence  the  ash-pit  is  always  cool  and  accessible, 
and  the  fire  is  kept  constantly  clean  and  effectually  grated.  About 
10%  of  the  fuel  drops  as  unburned  coke  with  the  ashes,  and  is 
carried  by  the  launder  to  a  concentrating  room  where  the  coke  is 
removed,  and,  since  it  still  retains  heat-value,  is  ground  with  flue- 
dust  and  fine  ore  and  made  into  briquettes  that  are  fed  to  the  blast- 
furnaces. When  the  Sterling  boilers  that  utilize  the  waste-heat  are 
to  be  repaired,  they  are  shut  off  by  damper,  and  the  furnace  gas 
is  turned  into  a  by-pass  flue  that  leads  directly  to  the  main  under- 
ground-flue to  the  main  stack.  The  draft  pressure  equals  l1/^  to 
2  in.  of  water. 

103.     CONVERTING  COPPER  MATTE. 

This  consists  in  treating  molten  matte  in  a  receptacle  lined 
with  refractory  material,  called  a  converter,  by  blowing  through 
it  compressed  air  which  burns  the  sulphur  of  the  matte  leaving 
molten  blister-copper  that  is  poured  from  the  converter  into  molds. 
In  making  steel  from  molten  pig-iron  the  same  process  is  used  but 
the  upright  converter  (See  Fig.  184),  is  employed.  In  treating 
copper  matte  the  barrel  converter  (Fig.  142),  which  revolves  about  a 
horizontal  axis,  is  used. 

104.     THE  COPPER  CONVERTER, 

Fig.  142  is  a  view  of  a  converter.  The  arrangement  and  position 
of  two  of  these  upon  the  floor  of  the  converter-house  is  shown  in 
Fig.  143  and  144.  In  Fig.  142  we  see  the  cylindrical  shell  and  the 
mechanism  by  which  it  is  rotated,  the  floor  being  broken  away  to 
show  the  hydraulic  operating-cylinder.  The  shell,  84  in.  diam.  by 
126  in.  long,  and  %  to  1  in.  thick,  is  made  of  the  plate-steel  in  two 
sections,  so  that  the  top  half,  parting  at  the  level  of  the  top  of 
the  end  spur-gear,  can  be  lifted  off,  leaving  the  lower  portion 
accessible  for  re-lining  with  ganister  (refractory  silicious  lining- 
material).  At  the  side  (not  shown  in  Fig.  142)  is  bolted  the  cast- 
iron  wind-box  which  is  of  the  length  of  the  shell  and  is  fitted  with 


OF    THE    COMMON    METALS. 


337 


14  tuyeres,  each  1  in.  diam.  To  one  head  of  the  shell  is  bolted  a 
cast-steel  open  gear-wheel  that  engages  the  upright  steel-toothed 
rack  connected  to  the  piston  of  the  hydraulic  cylinder  operated 
under  a  pressure  of  300  Ib.  per  sq.  in.  By  the  movement  of  this 


Fig.    142.      COPPER   CONVERTER. 


piston  and  rack  the  shell  is  tilted  into  any  desired  position,  for 
receiving  the  matte  or  for  pouring  out  slag  or  copper.  Riding 
rings  made  of  heavy  steel  rails  are  bolted  to  the  shell,  so  that  the 


338  THE    METALLURGY 

weight  is  borne  by  the  four  friction  rollers  on  which  it  revolves. 
The  friction  rollers  are  carried  by  heavy  cast-iron  base-frames.  TWTO 
inclined  tie-links  will  be  noticed  at  the  front  of  Fig.  142,  which  hold 
the  rack  meshed  in  gear.  When  the  shell  is  to  be  removed  for 
re-lining,  the  rack,  which  is  joined  at  the  head  of  the  piston  rod. 
is  swung  back  by  an  eccentric  at  the  foot  of  the  tie-links  so  that 
the  teeth  will  clear  those  of  the  spur- gear.  The  shell  is  lifted  by 
means  of  a  40-ton  traveling  crane  such  as  is  in  the  converter 
building,  Fig.  143.  In  Fig.  142,  close  to  the  top  of  the  cylinder,  is 
a  four-way  valve  controlled  by  a  hand-wheel  which  serves  to  tilt  or 
revolve  the  converter.  There  are  four  eyes  riveted  to  the  converter 
top  to  which  a  four-fold  lifting  chain  is  attached,  and  by  which 
the  top  can  be  removed  from  the  main  part  of  the  shell.  There  are 
also  four  eyes  riveted  to  the  main  part,  by  which  the  shell  can  be 
handled. 

105.     OPERATION  OF  CONVERTING. 

A  converter  stand  with  the  shell  is  constantly  in  use,  and  when 
the  lining  of  one  shell  is  consumed  the  converter  is  removed  and 
replaced  by  a  newly  lined  one.  After  lining,  the  converter  is  prepared 
as  follows.  Standing  upon  the  lining-floor  at  the  end  of  the  con- 
verter building  (See  Fig.  144),  it  is  dried.  The  interior  then  is 
brought  to  a  low-red  heat  by  means  of  a  wood  fire,  followed  by  one 
of  coke,  urged  by  a  blast  admitted  at  the  tuyeres.  A  converter  of 
the  barrel  type  of  the  size  above  specified  lined  at  least  12  in.  thick 
with  ganister,  takes  as  the  initial  double-charge  5  tons,  and  as  a 
final  double  charge,  just  before  it  has  to  be  re-lined,  12  tons  of  matte. 

A  charge  for  a  converter  is  tapped  from  a  reverberatory  furnace, 
or  from  the  fore-hearth  or  settler  of  a  blast-furnace,  into  a  steel 
ladle  handled  by  an  electric  traveling  crane  (See  Fig.  143).  The 
converter  is  rotated  to  bring  the  spout  from  beneath  the  hood 
(shown  above  it  in  the  illustration)  into  pouring  position,  the 
molten  matte  is  tipped  in  from  the  ladle,  and  the  converter  is 
returned  to  the  blowing  position,  in  which  it  appears  in  the  figure, 
the  pressure  of  blast  being  meanwhile  applied. 

The  first  blow. — The  first  blow,  or  the  slag-forming  stage  of  the 
process,  begins  when  the  converter  has  received  the  charge  of  molten 
matte.  A  50%  matte  would  have  a  composition  as  follows;  Cu  50%, 
Fe  20%,  S  24%,  corresponding  to  Cu2S  and  FeS,  with  a  little 
Fe304.  The  Fe3O4  and  the  Cu2S  are  but  slightly  affected  during 
this  blow,  the  first  entering  the  slag,  the  second  the  matte.  The 


OF   THE    COMMON    METALS.  339 

air  acts  with  increasing  energy  upon  the  FeS,  burning  both  the 
iron  and  sulphur  as  shown  by  the  following  reaction : 
FeS  +  30  =  FeO  +  S02 
23,800  66,400     71,000  ==  +  113,400 

This  is  an  energetic  exothermic  reaction  which,  in  spite  of  the 
volume  of  air  blown  in,  sustains  and  even  raises  the  temperature 
of  the  molten  matte.  The  S02  passes  off  in  the  escaping  gas,  and  the 
FeO,  with  other  bases,  takes  silica  from  the  lining,  which  accord- 
ingly is  rapidly  eroded.  The  reaction  is  at  first  slow,  but  it  rapidly 
increases  so  that  at  the  end  of  about  45  minutes  the  iron  sulphide 
is  burned  and  the  iron  slagged.  The  end  of  the  stage  is  determined 
by  the  appearance  of  the  issuing  flame,  the  first  greenish  border  of 
which  during  the  blow  changes  to  a  pale  permanent  blue  as  the  iron 
oxidizes.  Pieces  of  matte  are  thrown  into  the  converter,  and  by 
their  oxidation  make  the  charge  hotter.  If  too  hot,  sweepings  from 
about  the  converter  that  are  rich  in  copper  are  added  and  serve 
to  cool  it.  The  slag,  formed  in  result  of  the  addition,  increases 
the  total  volume.  The  converter  vessel  is  now  turned  into  pouring 
position,  the  blast  being  shut  off,  and  the  slag  is  poured  into  a 
steel  ladle  on  the  floor  placed  to  receive  it.  As  the  slag  flows  from 
the  converter  'the  skimmer'  passes  a  rabble  through  it  from  time 
to  time.  He  can  tell  when  the  matte  begins  to  escape,  and  signals 
for  the  vessel  to  be  returned  to  blowing  position,  whereupon  the 
hlast  is  turned  on  for  the  second  blow.  The  slag  from  the  first 
blow,  containing  1.5  to  2%  copper  and  about  0.5  to  1  oz.  silver  per 
ton,  is  sent  back  to  the  blast-furnace  and  smelted  to  extract  the 
remaining  silver  and  copper. 

The  second  blow. — At  the  beginning  of  this  blow  there  is  little 
iron  left,  and  the  copper  has  been  brought  to  the  stage  of  white- 
metal  of  75%  copper.  In  this  stage  the  oxidation  of  the  sulphur 
retained  by  the  copper  proceeds,  so  that  we  have : 

Cu2S  +  3O  =  Cu2O  +  S02 

The  Cu20  thus  formed,  reacts  on  the  CuS  not  then  burned  and 
produces  copper  with  the  escape  of  S02  gas  as  follows : 

Cu2S  +  2Cu20  ==  6Cu  +  SO2 

So  far  as  the  agitation  by  the  blast  permits,  the  molten  content 
of  the  vessel  separates  in  layers.  An  increasing  layer  of  copper  is 
at  the  bottom,  a  decreasing  layer  of  matte  is  above  it  and  slag  is 
on  the  top.  The  blast,  entering  horizontally  at  the  side,  6  to  12 
in.  above  the  bottom  blows  through  the  matte  that  is  floating  above 
the  copper  that  has  formed. 


340  THE    METALLURGY 

Under  normal  conditions  the  escaping  flame  is  of  white  gradually 
changing  to  a  rose-red  and  finally  to  a  brownish  red.  It  decreases 
in  length  and  volume  until,  at  the  last,  only  a  brick-red  flickering 
is  to  be  seen  in  the  escaping  gas.  The  determination  of  the  com- 
pletion of  the  operation  requires  care  and  experience.  If  carried 
too  far,  over-blown  copper  results. 

As  soon  as  the  content  of  the  vessel  has  been  changed  to  copper, 
the  converter  is  turned  down  into  pouring  position,  the  blast  being 
at  the  same  time  shut  off.  The  copper  is  poured  into  molds  that  are 
arranged  on  the  carriage  which  runs  on  a  bullion  track  beneath  the 
converter.  The  molds  hold  250  to  300  Ib.  of  copper  each,  and  there 
may  be  two  carriages,  each  carrying  15  to  20  molds.  The  converter 
is  turned  back  into  receiving  position  and  the  next  charge  is 
promptly  supplied. 

The  first  charge  of  a  newly  lined  converter  is  necessarily  small, 
but,  as  the  lining  becomes  eaten  out,  the  capacity  becomes  greater 
for  succeeding  charges.  In  converting  matte  of  50%  copper  the 
process  as  above  outlined  is  simple,  but  for  low-grade  matte  of  35 
to  40%  copper,  the  method  is  modified  by  'doubling',  as  it  is  called. 
Thus  into  a  newly  lined  converter  a  charge  of  2  tons  of  matte  is 
introduced.  This  is  blown  to  white-metal  in  the  first  stage.  The 
converter  is  turned  down,  the  slag  is  poured  off,  and,  upon  tilting 
the  converter  to  the  receiving  position,  3  tons  more  matte  is  added, 
making  a  double  charge  of  5  tons  in  all.  This  is  blown,  the 
slag  removed,  and  with  another  blow  it  is  finished  to  blister  copper. 
The  advantage  of  the  process  lies  in  the  fact  that  the  output  is 
increased,  and  that  the  white-metal  and  blister  copper  fill  the 
converter  and  make  a  suificient  quantity  of  blister  copper  fo(r 
pouring. 

The  final  charge  is  12  tons,  and  skice  large  charges  are  most 
profitable,  there  is  a  temptation  to  press  operations  to  the  limit. 
Toward  the  last,  the  lining  becomes  dangerously  thin,  showing  a 
red-hot  area  on  the  shell.  Sometimes  the  place  can  be  kept  cool  by 
a  stream  of  water  from  a  hose  and  the  charge  can  be  finished,  but 
if  not,  it  is  poured  into  a  ladle  and  transferred  to  another  converter. 

The  time  taken  for  converting  copper  matte  averages  40  to  50 
minutes  for  the  first,  and  50  minutes  for  the  second  blow.  For 
charging,  skimming,  and  poling  20  minutes  are  needed,  so  we  may 
say  that  it  takes  2  hours  for  the  cycle  of  operations  finishing  a 
charge.  A  first  charge  takes  little  more  than  an  hour,  and  a  final 
charge  more  than  two  hours. 

A  converter  lasts,  before  re-lining,  from  5  to  9  charges,  according 


OF    THE    COMMON    METALS.  341 

to  the  grade  of  the  matte  and  the  durability  of  the  lining.  Low- 
grade  matte,  which  contains  relatively  much  iron  sulphide,  corrodes 
the  lining  rapidly.  The  blast  pressure  varies  from  10  to  15  Ib.  to  the 
square  inch. 

106.     RE-LINING  CONVERTERS. 

When  a  converter  has  become  exhausted  it  is  removed  from  the 
stand  and  taken  by  the  crane  to  the  lining  floor.  The  top  is  unbolted 
and  removed  exposing  the  interior.  In  the  old  way,  the  interior 
of  the  vessel  was  cooled  with  a  stream  of  water  from  a  hose.  The 
objection  to  the  method  is  that  the  water,  taking  copper  salts  into 
solution,  has  a  corrosive  action  on  the  rails  beneath  the  converter 
that  frequently  have  been  wet  by  water  spilled  or  escaping 
containing  copper  in  solution.  In  the  new  way,  a  large  supply  of 
extra  shells  is  provided,  so  that  more  time  can  be  taken  for  cooling. 
The  lower  section  or  body,  laid  on  the  side  by  means  of  the  crane, 
is  cleaned  from  adhering  matte  and  slag  to  give  the  new  lining 
clean  surface  upon  which  to  adhere.  The  body  is  then  set  upright 
and  a  bottom  lining  of  ganister  (85%  quartz  and  15%  clay)  is 
tamped  in  at  a  level  6  in.  below  the  tuyeres.  A  sectional  prismoidal 
form,  or  pattern,  the  shape  of  the  intended  cavity,  is  set  in  position, 
and  between  this  and  the  shell  the  lining  is  firmly  tamped  with 
iron  tampers  or  rammers  to  the  level  of  the  top  of  the  form.  The 
form  is  removed  in  sections,  and  the  tuyere  openings  are  made 
with  a  pointed  rod.  The  top  section  is  lined,  and  put  on  and  bolted. 
The  joint  between  the  top  and  bottom  sections  is  made  with  a  softer 
mixture  than  the  rest  that  is  composed  of  72%  quartz  and  28% 
clay.  The  lined  shell  is  dried  and  heated  with  a  wood  followed  by  a 
coke  fire  as  has  been  described.  It  is  important  to  have  abundant 
converter-shells,  so  that  they  may  have  time  to  dry  thoroughly 
and  become  heated  before  using,  since  otherwise  the  life  of  the 
lining  is  shortened.  When  thoroughly  hot  the  shell  is  ready  to  be 
put  on  the  stand. 

Necessity  of  silicious  lining. — Ferrous  oxide  is  produced  in  the 
first  blowing,  and  by  itself  it  would  be  infusible.  To  slag  it,  silica 
must  be  supplied  with  which  it  can  combine.  Without  silica,  an 
infusible  mass  of  iron  oxide  would  accumulate  in  the  converter.  It 
is  not  possible  for  example,  to  use  a  converter  with  a  basic-lined 
or  water-jacketed  shell.  It  has  been  tried,  and  it  would  save  the 
time  and  expense  of  putting  in  linings  daily,  but  the  experiments 
have  failed.  The  life  of  the  lining  is  shorter  the  lower  the  grade 
of  the  matte,  because  with  much  FeQ  coming  from  the  matte,  a  large 


342  THE    METALLURGY 

quantity  of  slag  of  basic  quality  is  formed.  The  lining  also  is  liable 
to  fall  off  in  _umps  after  being  undercut  in  converting.  The  slag- 
produced  in  converting  60%  copper  matte  may  be  as  high  as  40% 
silica,  while  with  a  35%  matte,  the  resultant  slag  may  be  as  low 
as  25  per  cent. 

107.     SMELTER  AND   CONVERTER-PLANT. 

Fig.  143  and  144  represent  respectively  the  elevations  and  plan 
of  a  blast-furnace  plant  with  a  two-stand  converter-plant  attached, 
the  size  of  the  latter  being  indicated  by  the  number  of  stands  or 
stalls  in  which  copper  matte  can  be  blown.     In  the  elevation  the 
receiving  track   for   coke   is   shown   at   the   extreme   right   of   the 
illustration.     The  ore-bins  with  the  inclined  bottoms  are  shown  to 
be    on   the   same   level.     In   the    furnace   building    there    are    two 
matting  blast-furnaces,  each  42  by  144  in.  at  the  tuyere-level  an 
each  having  a  settler  or  fore-hearth  10  ft.  diam.    Blast  is  furnishe 
to  the  furnaces  from  a  power  house,  not  shown. 

In  Fig.  143  is  shown  the  semi-elliptical  flue  3  ft.  wide  by  7  f 
high,  leading  from  the  blast-furnaces.  This  crosses  the  end  of  th 
furnace  building,  as  indicated  by  the  dotted  lines  in  Fig.  144. 
is  connected  to  a  dust-chamber  which  leads  to  a  stack.  The  slag 
taken  away  over  an  electric-trolley  system  (See  Fig.  137)  enterin 
the  building.  The  slag-cars  are  brought  close  to  the  settlers 
receive  the  flowing  slag,  while  the  matte,  as  needed,  is  tapped  froi 
a  lower  tap-hole  into  the  steel  ladle  for  transferring  to  th 
converters.  A  platform  elevator,  at  the  end  of  the  furnace  buildin. 
elevates  slag  and  other  material  from  the  floor  of  the  converts 
building  to  the  charge-floor  for  the  blast-furnace. 

The  converter  building  has  at  one  end  the  lining  floor,  and 
commanded,  from  end  to  end,  by  a  40-ton  electric  traveling  cran 
which  serves  to  handle  the  converters,  to  supply  them  with  matt 
to  take  away  the  slag,  and  to  handle  all  materials  for  and  from  th 
converters.  For  each  stand  there  should  be  at  least  two  extra  shell 
or  six  in  all.  The  escaping  gases  from  the  converters  are  receive 
in  a  hood  attached  to  a  dust-chamber  (See  Fig.  143)  so  that  th 
particles  of  matte  blown  out  by  the  blast  are  collected.  From  th 
hood  a  flue  connects  with  a  dust-chamber  and  this  communicate 
with  a  stack. 

At  one  end  of  the  furnace  building  (See  Fig.  144)  is  the  mi] 
where  the  lining  material  is  prepared  for  the  converter.  It  consist 
of  a  10  by  16  in.  Blake  crusher  and  27  by  14  in.  rolls  for  coarsel 
crushed  quartz  needed  for  the  lining.  The  quartz  thus  coarsel 


OF    THE    COMMON   METALS. 


343 


crushed,  is  mixed  in  a  5-ft.  mud  mill  or  wet  pan  (Fig.  145)  with 
clay,  to  form  a  coherent  lining  material  or  ganister.  It  may  consist 
of  a  silicious  gold  or  silver-bearing  ore  mixed  with  clay.  The 
advantage  in  the  use  of  ore  is  that  the  gold  and  silver  are  taken  up 
by  the  copper  as  the  lining  is  consumed.  Compressed  air  from  the 


344 


THE    METALLURGY 


power  house  at  a  pressure  of  8  to  15  Ib.  per  sq.  in.  is  brought  by 
the  converter  blast  pipe  and  supplied  by  a  horizontal  blowing  engine 


OF    THE    COMMON    METALS. 


345 


capable  of  delivering  300  cu.  ft.  of  air  per  tuyere,  or  2500  to  3000 
cu.  ft.  per  min.  for  each  converter.  This  engine  of  the  duplex 
compound  type  has  steam  cylinders  17  and  32  in.  diam.  respectively, 
and  air  cylinders  36  in.  diam..  all  having  a  stroke  of  30  in.  and 
delivering  5600  cu.  ft.  of  air  per  minute  at  a  pressure  of  10  Ib. 


346 


THE    METALLURGY 


per  sq.  in.  and  capable  of  supplying  the  two  converters.  The  con- 
verters are  operated  with  a  hydraulic  cylinder  having  a  pressure 
of  175  to  500  Ib.  per  sq.  in.  from  a  pump.  The  demand  for  the  water 
is  irregular,  and  to  regulate  the  supply  the  hydraulic  accumulator 
(Fig.  147)  is  used.  It  consists  of  a  fixed  vertical  cylinder  in  which 
a  weighed  plunger  slides,  the  weight  being  a  sheet-steel  cylinder 


OF    THE    COMMON    METALS. 


347 


Fig.    147.      HYDRAULIC  ACCUMULATOR. 


348  THE    METALLURGY 

filled  with  pig-iron  weighing  15  to  20  tons.  When  water  is  used 
quickly,  the  plunger  falls  and  maintains  the  full  pressure,  while 
in  the  intervals  of  rest,  a  duplex  pump  fills  the  accumulator.  Slag 
is  received  from  the  converter  in  cast-steel  ladles  (Fig.  148),  and 
matte  is  supplied  from  the  settlers  of  the  blast-furnace,  or  from  the 
reverberatory  furnace.  The  ladles  have  a  capacity  of  5000  to  12,000 
Ib.  of  matte.  They  are  plastered  on  the  inside  with  a  coating  of 
clayey  loam  which  is  dried  before  using. 

The  ingots  or  pigs  of  copper  cast  in  the  molds  are  rough  on  the 
upper  surfaces,  owing  to  the  escape  of  occluded  gas  at  the  time  of 
solidification.  Blisters  form  on  the  surface,  hence  the  term  blister- 


Fig.  148.     CAST-STEEL  LADLE. 

copper.  For  electrolytic  refining  the  pigs  are  smelted  at  the 
refinery,  poled,  and  cast  into  anodes.  The  re-casting  gives  a  smooth 
and  even  surface  that  is  important  in  the  electrolytic  bath.  To 
obtain  smooth  anodes  at  the  United  Verde  works,  Jerome,  Arizona, 
a  reverbatory  melting  furnace  has  been  used.  Into  this  the  copper 
is  poured  from  a  ladle  as  the  converters  supply  it,  and  every 
morning  the  molten  copper  accumulated  is  poled,  to  reduce  the 
copper  oxide  present,  and  then  molded  into  anodes.  While  the 
poling  and  casting  is  going  on,  no  copper  can  be  put  into  the  melting 
furnace.  It  is  cast  into  ingots,  as  in  the  older  way,  and  when  the 
casting  of  anodes-  is  finished  and  the  furnace  is  empty,  the  ingots 
cast  during  the  interval  are  placed  in  the  melting  furnace  for 
beginning  the  next  charge. 


OF   THE    COMMON   METALS.  349 

108.  LOSS  IN  CONVERTING. 

The  gas  from  the  converter  consists  of  nitrogen  with  sulphur 
dioxide,  and  traces  of  volatilized  metals  (As,  Sb,  Te,  Pb,  Zn,  Cu,  and 
Ag).  The  loss  of  gold,  as  in  pyritic  smelting,  is  small,  but  the 
loss  of  silver  depends  upon  the  amount  of  the  volatile  metals.  A 
part  of  the  copper  and  silver  is  recovered  in  the  flues  with  the 
flue-dust  caused  by  the  blast.  The  silver  in  the  flue-dust  in  one 
case  was  found  to  be  28  to  64  oz.  at  the  branch  from  the  converter 
hood,  22  to  65  oz.  at  the  first  part  of  the  flue-chambers,  19  to  46 
oz.  farther  along  the  flue,  and  17.5  oz.  per  ton  near  the  stack.  Thus, 
the  larger  the  chamber,  the  greater  is  the  amount  recovered. 

The  loss  in  treatment  is  1  to  1.5  of  the  copper  and  2  to  2.5% 
of  the  silver.  In  attempting  to  treat  matte  from  silver-lead  blast- 
furnaces of  40%  copper  and  10.2%  lead,  the  silver  loss  was  serious, 
amounting  to  33  to  40%,  due  to  the  lead,  which  entirely  volatilized, 
and  carried  silver  as  it  does  when  present  in  a  copper-matting 
charge  in  a  blast-furnace. 

109.  COST  OF  CONVERTING. 

Hixon,  in  his  'Notes  on  Lead  and  Copper  Smelting  and  Refining,' 
gives  the  following  as  the  cost  per  pound  of  copper  of  converting 
matte  of  40  to  50%  copper  when  the  matte  is  re-melted  in  a  cupola 
before  converting. 

Cent. 

Re-melting  matte 0.20 

Labor  and  materials  for  re-lining 0.25 

Labor  on  converter 0.10 

Re-smelting  converter  slag 0.05 

Supplies    0.05 


Total  0.65 

For  direct-matte,  taken  molten  from  the  furnace,  the  cost  is  0.40c.  to 
0.45c.  or  $8  to  $9  per  ton. 

110.     THE  HYDRO-METALLURGY  OF  COPPER. 

The  wet  methods  of  extracting  copper  from  the  ore  consist  in 
obtaining  the  copper  in  water-solution  either  with  or  without  the 
aid  of  other  solvents.  The  copper  must  be  in  combination  with 
elements  that  cause  it  to  be  soluble  in  the  solvents  used.  From  the 
clear  or  filtered  solution  the  dissolved  copper  is  precipitated,  and 
then  melted  and  refined  by  igneous  methods.  In  order  to  extract 


350  THE    METALLURGY 

metal,  the  ore  is  first  rendered  suitable  in  form  for  dissolution,  by 
crushing,  and  if  necessary,  by  roasting  it, 

•  Copper  is  extracted  profitably  from  suitable  ore  of  as  low 
grade  as  0.5  to  1.5%  copper,  when  the  conditions  of  abundant  and 
easily  exploited  supply  and  cheap  labor  prevail.  Ore  containing 
the  copper  as  oxide,  carbonate,  or  sulphate  is  best  suited  to  the 
work,  but  ore  containing  lime,  magnesia,  ferrous  oxide,  or 
manganese  oxide,  is  less  desirable.  If  the  ore  contains  limestone, 
and  can  stand  the  expense  of  roasting  in  a  kiln,  then  the  caustic 
lime  can  be  leached  out  with  water  and  the  ore  treated  for  the 
extraction  of  the  copper.  Copper-bearing  sulphide  also  is 
profitably  treated  for  the  extraction  of  the  copper,  the  ore  being 
oxidized  until  the  copper  is  in  the  form  of  sulphate,  soluble  in 
water.  Sulphide  ores  also  have  been  roasted  with  salt,  to  bring 
the  copper  into  the  form  of  chloride,  which  then  is  extracted  with 
a  brine  solution. 

It  would  seem  that  extraction  methods  should  be  superior  to 
others  since  the  worthless  gangue,  which  is  the  largest  constituent 
of  the  ore,  remains  untouched,  and  the  solvent  acts  on  the 
relatively  small  quantity  of  valuable  metal.  Particularly  does  this 
seem  true  of  silicious  ore,  which  is  expensive  to  smelt,  and  the 
silica  of  which  in  no  way  interferes  with  leaching.  Experience, 
however,  has  taught  that  the  wet-treatment  of  ore  on  a  commercial 
scale  is  not  generally  profitable.  In  the  United  States,  the  ore, 
the  capital,  and  the  skill  in  handling  have  not  thus  far  resulted 
in  any  high  degree  of  success  with  wet-processes,  though  it  is 
possible  that  it  eventually  may  be  brought  about.  In  Mexico  the 
labor  conditions  and  climate  are  more  favorable.  With  large  and 
suitable  bodies  of  ore  it  is  possible  that  leaching  can  be  made 
profitable.  Such  work,  however,  must  be  conducted  under  the 
direction  of  skilled  and  experienced  metallurgists. 

111.     EXTRACTION    OF    COPPER    FROM    COPPER-BEARING 

SULPHIDE. 

We  may  divide  the  sulphide  process  into  the  two  following  types : 

(1)  Conversion   of   the   copper   into   soluble    form    (sulphate), 
leaching  the   copper  salt,   and   precipitating  the   copper  from   the 
filtrate  with  scrap-iron   (Rio  Tinto  process). 

(2)  Conversion   of  the   copper   into   chloride,    extraction   with 
a  salt-solution,  and  precipitation  with  scrap-iron  or  sulphur  dioxide 
(the  old  and  new  Hunt  and  Douglas  processes). 


OF    THE    COMMON    METALS.  351 

112.     THE  RIO  TINTO  PROCESS. 

This  well  known  method,  practised  so  successfully  at  Rio  Tinto 
in  Spain,  consists  in  preparing  large  heaps  of  copper-bearing  pyrite, 
allowing  it  to  oxidize  by  the  use  of  a  regulated  quantity  of  air  and 
water,  and,  when  the  copper  has  become  sulphate,  leaching  with 
water  to  extract  it.  The  clear  solution  is  conducted  to  tanks  filled 
with  pig-iron  where  the  copper  precipitates. 

When  the  copper  in  the  ore  occurs  as  chalcopyrite,  CuFeS2, 
or  as  covellite  CuS,  oxidation  proceeds  slowly  and  imperfectly,  and 
successful  working  requires  chalcocite  (copper  glance),  Cu2S.  It 
is  because  of  the  extent  of  the  orebodies  and  the  cheapness  of  labor 
that  the  method  has  been  successful. 

Method  of  working. — A  site  is  chosen  upon  impervious  sloping 
ground  for  suitably  draining  the  solution.  A  clayey  or  rocky 
bottom  is  required,  or  one  properly  puddled  and  lined  with  clay 
to  render  it  impenetrable. 

The  heaps  contain  100,000  tons  of  ore  and  are  constructed  as 
follows :  On  the  ground  is  first  arranged  a  net-work  of  flues,  12 
in.  square,  made  of  rough  stones.  Vertical  flues  or  chimneys  that 
connect  with  the  ground-flues  are  built,  50  ft.  apart,  as  the  heap 
i-s  made.  The  ore  is  broken  to  3  to  4  in.  diam.,  and  the  lumps  are 
screened  from  the  fine.  The  ore  is  dumped  and  spread  on  the  pile  in 
alternate  layers,  until  the  height  of  the  level  top  of  the  pile  is  30  ft. 
The  top  surface  is  formed  into  squares,  by  ridges  of  fine,  so  as 
to  insure  an  even  distribution  of  water  through  the  heap. 
Launders  are  provided  to  conduct  water  to  all  parts  of  this 
surface.  As  the  heap  is  forming,  water  is  supplied  to  extract  any 
copper  sulphate  already  existing.  Oxidation  starts  as  the  result 
of  the  wetting.  The  completed  and  wetted  heap  begins  to  oxidize 
rapidly  as  shown  by  the  heat  evolved,  the  temperature  of  the  air 
in  the  chimneys  rising  to  80°C.  As  the  temperature  rises,  the 
lower  ground-flues  are  closed  to  control  oxidation  and  to  distribute 
the  action  through  the  heap.  The  surface  assumes  a  brown  color, 
due  to  the  dehydration  of  the  basic  ferric  salt  that  forms,  and 
heating  is  made  apparent  by  this  drying  action.  Great  care  is 
taken  to  prevent  the  heap  from  catching  fire. 

By  the  combined  action  of  air  and  moisture,  the  following 
reactions  occur : 

(1)     FeS2  +  70  +  H20  =  FeSO4  +  H2SO4 
The  ferrus  sulphate  easily  oxidizes  to  ferric  sulphate  thus: 
(2)     2FeSO4  +  H2SO4  +  0  =  Pe2(SO4)R  +  H2O 


352  THE    METALLURGY 

The    ferric    sulphate    acts    on    the    chalcocite,    and    changes    it    to 
sulphate  thus : 

(3)     Fe2(S04)3  +  Cu2S  =  CuSO4  +  2FeSO<  +  CuS 
The  cupric  sulphide,  hitherto  unacted  upon  is  further  changed  as 
follows : 

(4)     Fe2(S04)3  +  CuS  +  3O  +  H20  =  CuS04  +  2FeS04  +  H2S04 
Reaction   (3)    is  relatively  rapid,   and  accordingly   about  half  the 
copper  goes  into  solution  in  a  few  months.     Reaction   (4)   is  slow 
but  in  two  years,  under  favorable   conditions  yields   80%    of  the 
remaining  half  of  the  copper. 

When  oxidation  has  advanced  as  far  as  is  safe,  water  is  applied 
at  the  rate  of  220  gal.  per  min.  until  the  soluble  copper  salts  are 
extracted.  The  flow  is  stopped  and  oxidation  is  resumed,  and  is 
followed  by  renewed  washing.  After  a  year  the  top  surface  needs 
're-tilling';  the  ridges  are  arranged  where  the  squares  formerly 
were,  and  the  launders  are  shifted  to  conform.  At  the  sides  of  the 
heap  the  ore  to  a  depth  of  a  few  feet  becomes  cemented  and  holds 
copper  salts.  The  sides  are  dug  down  in  terraces  to  expose  the 
copper  salts  and  extract  them  by  washing.  When  there  remains 
unextracted  but  0.3%  copper  in  the  ore,  extraction  is  considered 
complete. 

The  copper  sulphate  solution  that  flows  from  the  heap  contains 
ferric  sulphate,  and  to  prevent  it  from  consuming  iron  in  the 
precipitation  tanks,  the  ferric  sulphate  is  reduced  by  running  the 
solution  through  a  filter-bed  of  fresh  pyrite.  The  reaction  is 
as  follows: 

(5)  7Fe2(S04)3  +  FeS2  +  8H2O  =  15FeS04  +  8H2SO4 
The  bed  is  retained  within  a  reservoir  formed  by  a  masonry  dam 
across  a  small  ravine.  The  liquor  or  solution,  after  percolating 
the  bed,  remains  in  contact  until  drawn  into  the  precipitation-tanks. 
The  solution  entering  the  tanks  contains  Cu  0.4%,  Fe203  0.1%,  FeO 
0.2%,  H2S04  1.0%,  and  As  0.03%.  The  large  quantity  of  FeO  and 
H2SO4  is  due  to  the  fact  that  a  part  of  the  waste  liquor  from  the 
precipitation-tanks  is  pumped  back  arid  used  for  watering  the 
heaps,  so  that  the  solution  tends  to  concentrate. 

The  liquor  drawn  from  the  filter-bed  is  run  through  precipitation- 
tanks  containing  pig-iron  piled  in  open  order,  and  the  copper  is 
precipitated  (replacing  iron  which  dissolves)  in  the  form  of  a 
'cement  copper'  or  'copper  precipitate.'  The  tanks  are  arranged  on 
the  slope  of  a  hill,  and  the  liquor  passes  back  and  forth,  through 
the  tanks,  until  it  is  discharged  free  from  copper  from  the  lowest 
series.  Some  of  the  tanks  of  the  system  are  by-passed  daily,  the 


OF   THE    COMMON    METALS.  353 

liquor  meanwhile  going  through  the  remaining  ones.  The  tanks 
cut  out  are  drained,  and  all  the  pig-iron  is  removed  and  piled 
beside  them,  the  copper  attached  to  the  iron  being  meanwhile 
knocked  off  and  thrown  back  into  the  tank.  The  muddy  precipitate 
is  now  removed  to  the  cleaning  and  concentrating  plant,  while  the 
pig-iron  is  piled  back  into  the  tank  and  the  liquor  again  directed 
through  it.  Under  the  best  conditions  there  is  needed  1.4  tons  of 
pig-iron  per  ton  of  copper  precipitated. 

The  crude  precipitate  containing  10%  copper,  at  the  cleaning 
plant,  by  means  of  a  strong  jet  of  water  is  gradually  worked  over 
and  through  a  copper-plate  screen  placed  at  the  head  of  a  launder. 
The  over-size  of  the  screen,  consisting  of  leaf-copper  and  small 
pieces  of  iron,  is  thrown  into  a  heap  and  picked  over  by  girls  to 
remove  the  scrap-iron.  The  fine  that  passes  through  the  screen  is 
turned  over  under  a  stream  of  water  that  washes  out  the  dirt  and 
light  particles,  leaving  the  copper. 

At  the  head  of  the  launder  for  a  few  yards  is  found  No.  1 
precipitate,  94%  Cu  and  0.3.%  As.  Farther  along  is  No.  2  precipitate 
92%  Cu.  Next  comes  No.  3  precipitate,  which  is  fine  and  contains 
50%  Cu,  5%  As,  graphite  (from  the  pig-iron),  and  the  bismuth 
and  antimony  precipitated  from  the  liquor.  No.  1  and  No.  2 
precipitates  are  sacked  for  shipment,  and  No.  3  is  added  to  a 
blast-furnace  matting-charge,  the  copper  forming  matte,  while  the 
impurities  mostly  volatilize. 

113.     THE  OLD  HUNT  AND  DOUGLAS  PROCESS. 

In  this  process  the  copper-bearing  sulphide  is  crushed  and  then 
roasted  to  convert  the  copper  to  oxide.  The  copper  is  extracted  by 
a  combined  ferrous  chloride  and  salt-solution  as  cuprous  and  cupric 
chloride.  The  solution  is  filtered  and  the  copper  finally  is 
precipitated  upon  scrap-iron. 

The  copper-bearing  pyrite  ore  is  dry-crushed  to  a  4-mesh  size 
for  roasting.  It  is  then  thoroughly  roasted  until  the  copper 
sulphides  are  converted  to  oxide,  the  roasting  being  preferably  done 
in  one  of  the  mechanical  roasters  provided  with  a  fire-box  to  supply 
the  extra  heat  needed  to  effect  a  complete  roast.  The  cost  of  roast- 
ing is  the  principal  expense  of  the  process. 

The  roasted  ore  is  transferred  to  a  leaching  vat,  that  has  been 
carefully  painted  inside  with  asphalt  paint  to  protect  it  from  the 
action  of  the  solution.  ,  Jhe  percolating  solution  is  prepared  by 
dissolving  120  parts  Tor rou^r  chloride,  and  adrtitig  280  parts  green 
vitriol  (FeSO4  +  6H0O)  to  1000  parts  water.  ^R  the  solution 


354  THE    METALLURGY 


c 


fmmon  sait  is  added.     By  standing,  sodium  sulphate  crystallizes, 
leaving  a  strong  briny  solution  containing  ferrous  chloride.     When 


the  solution  meets  the  ore,  the  CuC^reacts^as  follows. 

(  1  )     3CuO  +  2FeCl2  =  Cu2Cl2  +  CuCL  +  Pe,Oa 
lublu.  ferric   oxido   and  the  soluble  cupric   chlorido,   if  prooont 

1P    nrp    anrl    f>nr>^RJnpd    i™    thn    mrnnnp    nf 


"til  IT'S  i 

(2)     3Cu20  +  2PeCl2  =  2Cu2Cl2  +  Pe208  +  2Cu 
But  the  copper  with  cupric  chloride  reacts  as  follows  : 

(3)     $Cu  +  CuCl2  =  Cu2012 
The  cuprous  chloride  is  soluble  in  the  salt-liquor. 

The  copper-containing  nitrate  from  the  filter-vats  passes  to 
precipitation  tanks  or  to  launders  containing  scrap  or  pig-iron,  and 
the  copper  precipitates  as  follows:  JL  Cu  Ci*  ^"  Cts^Cl  2. 


Ferrous  chloride  is  regenerated,  and  the  final  liquor,  after  passing 
through  the  boxes,  is  used  again. 

The  process  has  an  advantage  over  the  precipitation  of  copper 
sulphate  by  iron  in  that  less  iron  is  reo.uired.  In  reaction  (4), 
e»e-  equivalents  of  iron  {§j£)  precipitates  £mt  of  copper  (Ss&J.  The 
objection  urged  against  the  process  is  that  the  reaction  of  air  upon 
the  ferrous  chloride  solution  is  to  decompose  it,  forming  an 
oxychloride  thus: 

(5)     6FeCl2  +  3O  =SFe2Cle  +  Fe2O3 

The  ferric  oxide  formed  in  reaction  (1)  and  (2)  also  tends  to  clog 
the  filter.  To  avoid  the  difficulties  the  process  was  altered  to  the 
following  one. 

114.     THE  NEW  HUNT  AND  DOUGLAS  PROCESS. 

The  ore  is  crushed  and  roasted  as  described  for  the  old  process, 
then  treated  with  a  dilute  solution  of  sulphuric  acid  to  dissolve 
the  copper  oxide,  and  give  a  filtrate  of  copper  sulphate  containing 
ferrous  and  ferric  sulphates.  The  solution  is  delivered  into  a 
precipitating-tank,  and  a  solution  of  ferrous  chloride  is  added  which 
transforms  a  part  of  the  copper  sulphate  into  cupric  chloride 
(CuCl2). 

Sulphur  dioxide,  obtained  by  the  roasting  of  ore,  is  now  forced 
into  the  liquid.  It  precipitates  the  copper  as  an  insoluble  cuprous 
chloride,  as  follows  : 

(1)     CuCl2  +  CuS04  +  SO2  +  2H,O  =  Cu2012  +  2H2SO4 
The  cuprous  chloride  is  filtered  off  and  treated  with  milk-of-lime, 
or  with  iron  according  to  the  reaction  (4)  of  the  old  process.     The 


OF    THE    COMMON    METALS.  355 

consequent  FeCl2  is  used  in  the  treatment  of  the  next  charge,  while 
the  copper  precipitate  is  shipped  away.  The  sulphuric  acid  solution 
of  reaction  (1),  filtered  from  the  cuprous  chloride,  is  freed  from 
the  excess  of  S02  by  blowing  into  it  hot  air  from  an  injector,  and 
the  recuperated  solution  is  used  again  to  dissolve  ore. 

The  method  has  the  advantage  that  no  ferric  hydrate  is  formed 
to  clog  the  filter,  and  that  but  little  iron  is  needed  to  precipitate 
the  copper,  and  that  precipitated  copper  is  pure.  It  has  been  used 
for  the  extraction  of  copper  from  the  matte  rather  than  from  the 
ore. 

115.     EXTRACTION    OF    COPPER    FROM    OXIDIZED     ORES 

(NEILL  PROCESS). 

This  process  depends  upon  the  use  of  a  sulphur  dioxide  solution 
for  dissolving  the  copper.  The  copper  is  subsequently  precipitated 
by  heating  the  filtered  solution  and  expelling  the  S02  gas.  It  is 
used  preferably  upon  oxidized  ore,  such  as  native  carbonate  and 
oxide,  which  is  soluble  in  an  aqueous  solution  of  sulphur  dioxide,  but 
not  in  water.  The  process  can  be  used  also  in  the  treatment  of 
roasted  copper-bearing  ore.  Lime  and  magnesia  consume  sulphur 
dioxide  and  are  objectionable. 

For  oxidized  ore,  or  carbonate,  the  crushing  is  done  with  rolls, 
which  reduce  the  size  to  20-mesh.  The  ore  is  then  charged  into 
a  leaching  barrel,  like  that  used  in  chlorination,  and  water  is  added. 
A  stream  of  sulphur  dioxide  is  forced  into  the  barrel  by  means 
of  an  air-compressor  until  the  pulp  is  saturated  with  the  gas.  The 
saturation  is  maintained  some  hours,  until  the  copper  compounds 
have  dissolved.  The  sulphur  dioxide  is  produced  by  the  roasting 
of  iron  sulphide  in  a  pyrite-roaster. 

From  the  barrel  the  ore  passes  to  filter-presses  where  the 
solution  is  removed,  the  residual  tailing  being  rejected.  The 
solution  passes  to  the  precipitation  tanks  where  it  is  heated  by 
steam  and  boiled  until  the  SO2  gas  is  expelled. 

The  liquid,  now  freed  from  S02,  no  longer  retains  the  copper 
in  solution.  The  copper  comes  down  as  a  cupro-cupric  sulphite 
(CuSO3,Cu2S03  +  H20),  a  heavy  crystalline  compound  of  a  dark- 
red  color,  containing  49.1%  copper.  The  supernatant  solution  from 
the  precipitating  tank  runs  to  waste  through  launders  containing, 
as  a  precaution,  scrap-iron  to  insure  removing  the  last  of  the 
copper.  The  precipitated  sulphite  readily  settles  from  the  solution. 
It  is  washed  by  decantation,  dried,  and  reduced  and  melted  in  a 
reverberatory  furnace  giving  metallic  copper. 


356  THE    METALLURGY 

The  process  has  the  advantage  that  a  unit  of  copper,  converted 
into  cuprous  sulphite,  needs  but  half  the  sulphur  that  would  be 
required  to  convert  it  to  cupric  sulphate.  Cuprous  sulphite  is  here 
precipitated  from  the  solution  without  the  use  of  scrap-iron.  This 
is  an  advantage  in  remote  districts  where  the  cost  of  transportation 
is  high.  Sulphur  dioxide  has  little  action  upon  other  metals,  and 
thus  a  pure  copper  is  furnished  by  this  process.. 

116.     THE  HENDERSON  PROCESS. 

The  well  roasted  residue,  or  cinder,  resulting  from  the  pyrite 
used  in  making  sulphuric  acid,  contains  2  to  4%  copper,  with  silver 
and  gold.  All  these  metals  can  be  extracted  by  a  chloridizing 
roast  followed  by  leaching  with  weak  liquor  from  a  previous 
operation,  and  containing  water  and  dilute  hydrochloric  acid.  The 
copper  in  the  clear  filtrate  is  precipitated  upon  scrap-iron. 

Operation  of  plant. — Fig.  149  shows  the  plan  and  a  transverse 
sectional  elevation  of  a  200-ton  plant  of  the  Pennsylvania  Salt 
Manufacturing  Co.,  Natrona,  Pennsylvania. 

The  cinder  (red-roasted  or  burned  pyrite)  that  is  brought  from 
the  various  sulphuric  acid  plants  throughout  the  country  is  ground 
dry  to  20-mesh  in  a  pan-mill,  Fig.  145,  and  mixed  during  the  grind- 
ing with  12%  of  the  weight  of  salt. 

The  mixture  is  raised  by  a  belt-elevator  to  storage  bins  (not 
shown)  commanding  the  charge  floor  k.  It  is  weighed  in  5-ton 
charges  and  put  into  the  charge-tubes  of  the  muffle  roasters  shown 
in  the  longitudinal  sectional  elevation,  Fig.  150.  There  are  four 
charge-tubes,  each  20  in.  diam.  for  each  furnace.  The  gases  from 
the  fire  pass  along  the  14-in.  space  above  the  8  by  35-ft.  muffle- 
hearth  containing  the  ore,  thence  by  a  flue  downward  to  the  space 
below  the  muffle,  and  finally  by  a  main  underground-flue  to  the 
stack.  The  gases  from  the  roasting  ore  pass  by  an  18-in.  pipe 
to  condensing  towers  a,  filled  with  lump  coke  that  is  wet  by  a 
spray  of  water  above.  The  water,  coming  in  contact  with 
the  ascending  gas,  absorbs  the  chlorine  and  hydrochloric  acid. 

Raw  pyrite  is  charged  with  the  ground  cinder  to  make  the 
sulphur  content  I1/-?  times  that  of  copper.  The  charge  is  heated 
to  a  visible  red  heat  (525 °C.)  and  well  stirred  during  8  hours. 
When  finished,  it  is  drawn  out  upon  the  floor,  allowed  to  cool, 
shoveled  into  charge-cars,  raised  by  platform-elevator  to  the 
charge-floor  level,  and  put  into  the  leaching  tanks  d,  each  of  which 
is  12  by  14  ft.  in  size. 

The  ore  is  first  lixiviated  with  a  weak  liquor  from  a  previous 


OF    THE    COMMON    METALS. 


357 


operation  to  remove  most  of  the  copper.  The  solution  becomes  a 
strong  solution.  The  ore  is  then  treated  with  water,  to  remove  the 
remaining  copper  and  the  solution  becomes  the  weak  solution  of  the 
succeeding  operation.  Finally,  the  weak  solution  of  hydrochloric 
acid  from  the  towers  a  is  applied,  dissolving  the  cupric  oxide  and 
cuprous  chloride,  hitherto  insoluble.  The  residue,  called  'purple 


ii  HI  riin 


ini  nininiriinin  i  n  i 


a 


JUUUUi 


LJ 


•J/yacs    for 
Mi/fs,  Sfy/ns,  Bo/'/erj, 
C'naer  and  -Saff 


J£L 


149.      HENDERSON-PROCESS   PLANT. 

ore'  is  shoveled  from  the  vats  to  the  floor  c  and  thence  discharged 
into  the  railroad  cars  below. 

The  weak  solution  is  sent  to  the  lixiviation  tanks.  The  strong 
solution  when  the  specific  gravity  reaches  18 °B.  is  drawn  to  tanks 
12  by  12  by  6  ft.  where  the  copper  is  precipitated  upon  scrap-iron. 
The  tanks  have  false  bottoms  of  slats  2  ft.  above  the  bottom. 
Live  steam,  directed  into  the  solution,  agitates  it.  The  copper 
precipitating  upon  the  iron  works  down  between  the  slats  to  the 
bottom  of  the  tanks  and  is  removed  to  tanks  g,  10  by  10  by  5  ft. 
The  solution  from  this  tank  is  drawn  into  launders  containing 
scrap-iron  as  a  guard,  and  to  retain  any  remaining  particles  of 


358 


THE    METALLURGY 


precipitate.     The  precipitate  is  90%  copper,  35  oz.  silver,  and  0.15 
oz.  gold  per  ton.     It  is  sold  to  the  blue-vitriol  makers  who  pay 


95%  of  the  silver  and  the  full  value  of  the  copper  and  gold. 

The  cost  of  treatment  by  the  process,   with  common  labor  at 
$1.50  per  day,  is  $1.87  per  ton  of  cinder  treated. 


PART  VII.     LEAD 


PART  VII.  LEAD. 

117.  THE  LEAD  ORES. 

The  lead  ores  are  those  in  which  lead  is  the  principal  constituent. 
The  term  is  applied  also  to  mineral  aggregates  consisting  of  more 
than  10%  lead.  The  lead  ores  may  be  divided  into  two  classes: 
the  sulphide  and  the  oxidized.  The  terms  are  used  only  according 
to  the  constituent  that  is  in  excess,  in  many  lead  ores  both 
sulphides  and  oxides  are  found.  Ore  containing  no  lead  is  called 
dry,  and  when  carrying  lead,  leady.  The  latter  term  is  the  opposite 
of  dry,  but  we  do  not  term  a  leady  ore  a  wet  one. 

Galena. — Pure  galena  contains  86.6%  lead  and  13.4%  sulphur. 
In  nature  it  occurs  with  gangue  or  vein-matter.  When  there  is 
much  of  the  latter  it  can  readily  be  concentrated.  The  following 
table  gives  an  idea  of  the  lead-content  of  ore,  before  and  after 
dressing : 

GALENA  ORES. 

Raw  ore.  Concentrate. 

Pb,  Pb,  Ag,  oz. 

Locality.                                                                     %  %  per  ton. 

Minnie  Moore,  Wood  River,  Idaho 62.0  80.0 

Rockville,  Wisconsin ...  0.3 

St.  Joseph,  Missouri   7.0  70.0 

Kellogg,  Idaho 11.0  60.0  30.0 

Col.  Sellers,  Leadville,  Colorado 10.0  55.0  19.8 

Galena  from  the  Mississippi  Valley  contains  little  silver,  but 
from  the  Rocky  Mountain  region  it  is  not  only  argentiferous,  but 
may  contain  gold.  The  precious  metals  as  well  as  the  lead  determine 
the  value.  Metallic  sulphides,  such  as  pyrite  and  blende,  are  often 
associated  with  galena,  and  with  the  gangue  may  carry  so  much 
of  the  gold  and  silver  that  concentrating  leads  to  a  serious  loss 
of  the  metals  and  is  omitted.  If  by  hand-picking  ore  can  be  brought 
to  contain  30  to  40%  lead,  it  is  a  desirable  ore  for  the  smelter.  When 
of  this  tenor  in  lead,  and  free  from  other  sulphides,  it  carries  but 
5%  sulphur  and  needs  no  preliminary  roasting,  and  is  smelted 
directly. 

Oxidized  lead  ores. — Little  lead  oxide  is  found  in  nature.  The 
ores  classed  here  under  oxidized  ores  are  the  result  of  the  alteration 


362  THE    METALLURGY 

of  galena.  They  include  the  carbonate  (cerussite)  and  the  sulphate 
(anglesite)  of  lead.  The  minerals  are  mixed  with  metallic  oxides 
and  vein  matter  or  gangue  in  nature,  and  when  sandy  or  earthy,  the 
ore  is  called  sand  or  soft  carbonate,  and  when  hard  and  stony, 
hard  carbonate.  In  many  deposits  we  find  ore,  that  originally  was 
galena,  profoundly  altered  to  cerussite  or  anglesite.  The  subjoined 
table  gives  the  composition  of  some  of  the  so-called  carbonates. 

CARBONATE   ORES. 


Locality. 

Pb, 

% 
72  0 

SiOL>, 

% 

Fe, 

% 

CaO, 

% 

s, 

% 

Ag.  oz. 
per  ton. 

38  0 

25  0 

Leadville    Colorado    

.      21.0 

22  5 

18.2 

2.4 

0.9 

65.0 

Red  Mountain    Colorado 

18  4 

41.6 

11.4 

1.7 

1.8 

128.0 

33.2 

3.0 

24.1 

1.1 

2.0 

27.5 

Bingham     Utah                      

.      51.5 

12.5 

2.6 

3.2 

6.0 

21.1 

Horn  Silver  mine,   Frisco,  Utah.. 

.      50.0 

15.2 

3.4 

0.5 

8.3 

78.3 

Of  the  ores  of  the  table,  that  from  Eureka,  Nevada,  contains 
4.2%  of  arsenic,  which  forms  an  arsenical  speiss  when  smelted. 
The  Horn  Silver  ore,  apparently  oxidized,  has  the  lead  in  the 
form  of  anglesite  (PbSOJ,  and  matte  is  formed  from  it  in  smelting. 
In  oxidized  ores  the  silver  is  apt  to  occur  as  a  chloride,  the  gold 
probably  is  native. 

There  are  many  lead  minerals  but  those  not  mentioned  occur 
in  small  quantity  and  are  not  considered  among  the  commercial 
lead  ores. 

118.     RECEIVING,  SAMPLING,  AND  BEDDING  ORES. 

Where  ore  is  treated  in  a  small  way  for  the  recovery  of  the 
lead,  as  in  Missouri,  no  particular  provision  is  made  for  storage. 
In  various  smelting  works  in  the  Rocky  Mountain  region,  where 
lead  ores  are  treated  with  others  by  methods  of  silver-lead 
smelting,  and  where  ores  are  bought  outright  for  treatment,  the 
handling  becomes  complicated.  Such  plants  are  called  custom 
works.  A  plant  treating  ore  from  a  single  mine  is  called  a  mine's 
works,  and  here  less  attention  is  given  the  sampling  and  storing 
of  the  ore.  In  customs  works,  therefore,  ores  of  many  kinds  are 
received,  some  containing  lead,  some  having  little  lead  but  carrying 
silver  and  gold. 

The  ore  is  received  in  lots  of  a  few  tons  up  to  those  of  several 
carloads.  Each  lot  is  separately  weighed,  sampled  as  fully  described 
in  the  chapter  on  sampling,  assayed,  and  purchased.  The  company 
then  is  free  to  treat  the  ore  as  it  pleases.  If  different  kinds  of 
ore  were  smelted  separately  the  process  would  involve  endless 


OF  THE  COMMON  METALS. 


363 


change  and  labor,  and  so  it  has  become  the  custom  to  'bed'  the 
ore  in  large  bins  holding  several  hundred  tons.  When  so  bedded 
the  mixture  is  treated  as  a  single  ore.  The  different  kinds  of  ore 
are  unloaded  separately  into  the  bin  and  each  kind  is  spread  out 
in  an  even  layer  before  the  succeeding  one  is  added.  This  is 
indicated  in  Fig.  151,  and  it  is  seen  that  we  thus  have  a  series 
of  Ia3rers  in  a  bin.  When  the  ore  is  to  be  used,  shoveling  is  done 
at  the  floor  and  all  parts  above  fall  down  and  mix,  since  a  steep 
face  of  ore  is  constantly  maintained.  Thus  a  uniform  mixture  of  the 
different  ores  is  obtained  for  smelting,  and  so  long  as  we  are  using 


t  : 


Fig".    151.      ORE-BED. 


ore  from  this  bin,  the  quality  remains  constant.  The  supply  remains 
practically  unchanged  in  quality  several  days,  and  the  content  of 
the  bed  is  treated  in  the  books  of  the  company  as  a  single  ore. 

In  the  laboratory  the  aggregate  analysis  of  the  bed  is  obtained 
as  follows:  A  list  of  the  ores  and  the  dry  weights  is  prepared. 
The  chemist  weighs  out,  on  his  balance,  from  the  reserved  samples 
of  each  of  the  ores,  an  amount  proportionate  to  the  weight  of  ore 
in  the  bin.  The  total  portion,  amounting  to  one  or  two  ounces,  is 
thoroughly  mixed,  and  from  it  portions  are  taken  for  the 
determination  of  Si02,  Fe,  CaO,  and  S.  The  mixture  may  also  be 
assayed  for  silver,  gold,  and  lead,  as  a  check  on  the  calculated 


364  THE    METALLURGY 

content  obtained  from  the  weights  and  calculated  assay  of  the 
individual  ores.  The  determinations  thus  made  are  used  in  com- 
puting the  charge. 

Besides  ore  bedded  in  this  way,  lots  that  would  fill  a  large 
bin  remain  unmixed,  and  the  ore  is  treated  as  a  separate  item  of 
the  charge.  Small  lots,  of  which  a  moderate  amount  is  to  be  used 
upon  the  charge,  also  may  be  kept  separate,  and  such  are  called 
'side  ores.' 

Crushing  and  bedding  ores  for  roasting. — Sulphide  ore  that  is 
to  be  roasted  also  may  be  bedded.  When  thus  made  uniform,  it 
becomes  known  and  is  handled  better  at  the  roaster,  and  it  makes 
a  uniform  product  for  the  blast-furnace.  Such  beds  are  made  to 
contain  the  proportions  of  pyrite,  silica,  and  galena  that  work 
best  in  the  roaster. 

Crushing  the  sulphide  is  often  performed  in  two  stages.  The 
coarse  crushing  of  lumpy  ore  is  done  with  rock-breakers,  either  of 
the  jaw  or  the  gyratory  type,  the  ore  being  reduced  to  %-in.  size. 
It  is  then  passed  through  rolls,  36  in.  diam.  by  14  in.  face,  where 
it  is  crushed  to  pass  a  3  to  10-mesh  screen  according  to  the  nature 
of  the  ore.  A  pyritiferous  ore  need  be  crushed  no  finer  than  3-mesh. 
while  ore  carrying  galena  and  blende  roasts  better  when  crushed 
to  10-mesh. 

Sulphide  ore  is  preferably  bedded  before  roasting,  because  the 
men  then  soon  learn  how  best  to  roast  it,  whereas  with  constant 
changes,  they  fail  in  this.  Ores  of  different  composition  roast 
to  advantage  when  judiciously  mixed  with  a  silicious  pyrite  ore.  A 
pyrite  ore,  which  readily  starts  to  roast,  assists  the  slow  galena  or 
blende.  By  combining  different  kinds  we  obtain  a  mixture  that 
agglomerates  only  when  the  roasting  is  completed  and  the  roasted 
material  is  ready  to  be  withdrawn  from  the  furnace.  A  bed  formed 
of  10  to  15%  SiO2,  20  to  28%  Fe,  and  20  to  28%  Pb,  roasts  well. 
Mixtures  containing  less  lead  and  more  pyrite  than  this  roast 
readily,  but  ore  that  is  pulverulent,  when  roasted,  tends  to  make 
more  flue-dust  in  the  blast-furnace,  while  with  the  proportion  of 
lead  above  specified  it  tends  to  sinter  and  make  a  desirable  lumpy 
product  for  the  blast-furnace. 

119.  THE  SMELTING  OF  LEAD  ORES. 

When  lead-bearing  ores  are  to  be  smelted  only  for  the  lead 
content,  as  is  done  in  parts  of  the  Mississippi  Valley,  a  simple 
plant  with  a  reverberatory  furnace,  or  the  American  ore-hearth,  is 
sufficient.  In  the  Rocky  Mountain  region  the  lead  ore  is  not  smelted 


OF   THE   COMMON   METALS. 


365 


to  recover  only  the  lead.  The  lead  of  the  ore  is  employed  as  a 
collector  of  the  gold  and  silver  of  other  ores  that  are  smelted  at 
the  same  time.  In  the  first  case,  a  large  part  of  the  lead  is  recovered 
cheaply  and  simply,  but  a  part  is  lost  in  the  resultant  slag.  In 
silver-lead  smelting  it  is  essential  that  the  slag  be  comparatively 
free  from  lead  and  the  consequent  silver. 

120.     REVERBERATORY   LEAD    SMELTING. 

The  treatment  of  lead  ore  in  reverberatory  furnaces  has  not 
made  much  headway  in  the  United  States.  There  are  two  reasons 
for  this :  In  the  silver-lead  districts,  the  ore  has  not  been  of 
sufficient  grade  in  lead  to  warrant  the  treatment,  and  lead  ore  has 
been  in  great  demand  as  a  collector  to  mix  with  other  ores. 
Secondly,  in  the  Mississippi  Valley,  where  silver-free  high-grade 
lead  ores  occur,  the  question  of  skilled  labor  for  reverberatory- 
furnace  work  has  had  an  influence. 

The  reverberatory  lead  furnace. — Fig.  152  and  153  represent 
one  of  the  large  recent  furnaces,  16  by  9-ft.  hearth  dimensions, 


HORIZONTAL  SECTION  ON  LINE  G,  H 


Fig.    152.      LEAD-SMELTING  REVERBERATORY   FURNACE    (PLAN). 

having  a  fire-box  8  ft.  by  20  in.  or  of  14  sq.  ft.  grate-area.  The 
floor  or  bottom  of  the  hearth  slopes  from  the  fire-bridge  to  the 
corner  near  the  external  well  or  basin  /  at  the  cool  end  of  the 
furnace.  The  flame  passes  to  the  stack  (not  shown)  by  a  flue 
at  a.  The  charge  is  dropped  into  the  furnace,  as  needed,  through 
a  12-in.  hole  in  the  middle  of  the  roof.  There  are  four  working- 


366 


THE    METALLURGY 


doors  on  each  side,  so  that  the  interior  is  easily  reached  to  spread, 
rake,  or  withdraw  the  charge.  The  lead,  as  it  forms,  drains  to 
the  basin  /,  from  which  it  is  dipped  from  time  to  time  as  it 
accumulates,  and  molded  into  bars  or  ingots. 

Operation  of  the  furnace. — The  operation  is  divided  into  two 
stages:  the  first  that  of  roasting,  or  oxidation;  the  second, 
reduction. 

Oxidation. — Four  tons  of  ore,  crushed  to  5-mesh  size,  is  dropped 
into  the  furnace  from  the  hopper  and  spread  over  the  hearth  in 
a  layer  3  in.  deep.  An  oxidizing  fire  is  maintained  in  the  fire-box 
to  raise  the  temperature  of  the  charge  to  visible  red  (500  to  600°C.)- 
The  roasting  is  kept  up  three  or  four  hours,  and  continued  only  until 


Firebrick 


LONGITUDINAL  SECTION  ON  LINE  A,  B. 


Fig.   153.      LEAD-SMELTING  REVERBERATORY  FURNACE    (ELEVATION). 

an  incomplete  but  definite  degree  of  oxidation  has  resulted.  The 
reaction  is  as  follows  : 

(1)     2PbS  +  70  =  PbO  +  PbS04  +  SO, 

The  galena  is  converted  in  part  into  oxide,  in  part  into  sulphate. 
A  part  remains  as  sulphide.  The  temperature  is  kept  at  the  required 
degree  and  the  charge  is  frequently  raked  to  expose  new  surface  to 
the  action  of  the  air,  and  to  prevent  the  agglomeration  of  the 
charge. 

Reduction.—  The  fire-box  is  filled  with  a  thick.  bed  of  coal  to 
give  a  neutral  flame,  and  the  temperature  is  raised  to  the  point 
at  which^the  charge  begins  to  soften,  but  not  to  melt.  The  oxide  and 


of  lead  react  upon  the  unchanged  galena  thus  : 

(2)  PbS  +  2PbO  ===  3Pb  -f  S02 

(3)  PbS  +  PbS04  =  2Pb  -\3$<d2 

The  metallic  lead  begins  to  flow.  To  stiffen  the  charge  and  make 
it  less  fusible  and  more  open  in  texture,  slaked  lime  is  added  and 
stirred  in.  The  change  is  rabbled  at  intervals  to  promote  the 


OF    THE    COMMON    METALS.  367 

reactions.  Finally  the  flow  of.  lead  ceases,  but  the  pasty  residue  still 
retains  half  the  original  lead. 

Further  treatment. — To  extract  lead,  a  second  roasting  takes 
place,  followed  by  a  second  period  of  reaction.  It  takes  several 
of  these  operations  to  extract  all  the  lead  possible.  Toward  the 
end  there  remains  no  lead  sulphide  to  react  with  the  lead  sulphate 
and  oxide.  To  reduce  these,  slack  coal  or  charcoal-breeze  is  mixed 
in,  and  a  further  portion  is  obtained.  Each  successive  operation 
takes  less  time,  and  the  temperature  becomes  a  little  higher  as 
the  charge  becomes  stiff  er.  The  lead,  as  it  drains  away,  is  received 
in  the  basin  /,  and  after  skimming  is  molded  into  bars. 

The  residue,  after  extracting  all  the  lead  possible,  is  a  gray 
slag  still  containing  12  to  30%  lead,  and  having  one-fourth  the 
weight  of  the  original  charge.  It  may  be  sent  to  a  blast-furnace 
where  the  lead  can  be  recovered. 

The  process  requires  twelve  hours  for  a  four-ton  charge  and 
consumes  45%  or  1.8  ton  of  coal.  It  is  suited  only  to  concentrate 
and  to  ores  rich  in  lead,  but  not  to  those  of  more  than  4  to  5% 
silica.  Silica  forms  a  silicate  with  the  lead  oxide  and  carries  the 
lead  into  the  slag.  A  small  amount  of  metallic  sulphide,  such 
as  pyrite  or  blende,  is  not  harmful;  indeed  pyrite  is  beneficial  at 
'the  roasting  stage.  Limestone,  dolomite,  blende,  and  iron  oxide 
stiffen  the  charge  and  prevent  premature  melting. 

121.     THE  ORE-HEARTH. 

The  ore-hearth  cannot,  as  regards  capacity,  or  cost  per  ton 
of  ore  treated,  compete  with  the  reverberatory  furnace,  nor  can 
it  be  used  in  silver-lead  smelting.  As  compared  with  the 
reverberatory  it  can  be  quickly  started  or  stopped,  and  put  in 
operation  with  little  cost  for  fuel,  so  it  well  serves  the  purpose 
of  extracting  the  lead  from  small  amounts  of  low-silver  ore,  from 
time  to  time,  by  the  men  who  themselves  have  mined  the  ore. 

The  hearth. — !*ig.  154,  represents  a  sectional  elevation,  a  front 
elevation  of  the  lower  part,  and  a  plan  of  an  American  ore-hearth. 
It  consists  of  a  cast-iron  pan  or  crucible  a,  2  by  2%  ft.  by  1  ft. 
deep,  to  contain  a  bath  of  lead,  and  is  built  into  the  brick  work  q. 
The  back  p  and  the  two  sides  «,  n,  above  the  crucible,  are  water- 
cooled  castings.  The  blast  from  a  fan-blower  (not  shown)  enters 
by  the  tuyere-pipe  &  through  the  back  at  o.  At  g  is  a  sloping  cast- 
iron  plate  called  the  work-stone,  and  at  i,  a  pot  placed  to  recover 
the  lead  that  flows  down  over  the  work-stone.  The  pot  is  kept 


368 


THE    METALLURGY 


hot  by   a   wood   fire   below.      The    structure    is   surmounted   by   a 
brick  top  to  receive  and  carry  off  the  fumes. 

Operation. — By  means  of  the  blast,  a  glowing  coal  fire  is  made 
that  fills  the  hearth.  Eesidue  from  the  previous  run,  containing 
metallic  lead,  and  15  to  20  Ib.  galena,  not  finer  than  pea-size,  is 
spread  over  the  fire.  The  charge  soon  becomes  red-hot,  and  the 
lead,  set  free,  finds  its  way  to  the  crucible  at  the  bottom.  More 


VERTICAL  SECTION  ON  LINE  6,  D. 


FRONT  ELEVATION. 


HORIZONTAL  SECTION  ON  LINE  A,  B. 


Fig.    154.      AMERICAN  ORE-HEARTH. 

ore  is  then  added,  and  the  material  in  the  hearth  is  pried  up  gently 
with  a  bar  to  keep  the  mass  open  and  hot  throughout.  Lumps 
form,  and  are  drawn  out  on  the  work-stone,  and  gray  slag  that 
forms  at  the  same  time  is  separated  and  the  rich  residue  returned 
to  the  hearth.  Ore  and  fuel  are  again  added,  15  to  20  Ib.  at  a 
time,  and  operations  continue  until  lead  fills  the  crucible,  while  on 
the  top  floats  the  fuel,  unreduced  ore,  and  half-fused  material.  One 
man  with  a  bar  at  intervals  loosens  and  stirs  the  charge,  raising 


OF    THE    COMMON    METALS.  369 

it  slowly,  while  another  with  a  shovel  draws  upon  the  work-stone 
the  half-fused  mass  floating  on  the  lead.  Here  he  separates  and 
rejects  the  gray  slag,  and  returns  the  rich  residue  to  the  charge. 
A  fresh  charge  is  then  added,  and  the  work  progresses  in  the  manner 
described.  The  lead  overflows  the  crucible  and  runs  down  a  groove 
made  in  the  work-stone  into  the  kettle.  When  the  kettle  fills,  the 
lead  is  skimmed  and  ladled  into  molds. 

The  reactions  in  which  the  lead  is  reduced  are  like  those  of 
the  reverberatory  process.  In  the  ore-hearth  also,  the  glowing 
fuel,  acting  upon  the  lead  oxide,  reduces  it  to  metallic  lead. 

To  operate  the  ore-hearth,  a  blower  and  power  to  run  it  are 
needed.  Much  lead  is  volatilized,  and  so  the  treatment  is  not 
suited  to  argentiferous  galena.  The  gray  slag  that  is  produced 
still  consists  of  35  to  40%  lead,  and  is  sold  to  smelting-works.  The 
direct  recovery  of  the  lead  is  75  to  85%,  the  higher  figure  having 
been  obtained  in  recent  practice. 

122.     SILVER-LEAD    SMELTING. 

This  is  a  blast-furnace  method  of  treatment,  applicable  to  a 
great  variety  of  ores  containing  lead,  silver,  gold,  and  even  copper. 
By  it,  ore  containing  the  precious  metals  with  no  lead,  are  treated 
with  lead-bearing  ores,  thus  using  the  lead  of  one  ore  as  the  collector 
of  the  gold  and  silver  of  another.  The  method  is  the  most  effective 
means  of  doing  this.  The  precious  metals  are  extracted  from  the 
ore  by  a  blast-furnace  treatment,  using  lead-bearing  ores,  carbon- 
aceous fuel,  and  flux. 

Oxidized  ore  can  be  directly  smelted  in  the  blast-furnace,  but 
sulphide  ore  is  first  roasted.  Methods  of  roasting  are  described 
in  the  chapter  on  roasting,  and  in  the  chapter  under  'crushing  and 
bedding  of  ores  for  roasting. '  These  have  become  important 
preliminary  operations  for  the  treatment  of  ore  before  it  is  smelted. 

The  ore  to  be  treated  is  charged  into  the  blast-furnace  as  in 
iron  or  copper  smelting,  with  a  calculated  quantity  of  flux  which 
for  lead  ore  is  iron  ore  and  limestone.  The  precaution  is  taken 
to  use  lead-bearing  ore  to  make  the  lead  content  of  the  charge 
at  least  10%.  It  has  been  found  that  if  a  smaller  proportion  of 
lead  than  this  is  used,  the  precious  metals  are  not  so  well  collected 
in  the  base-bullion,  or  work-lead,  produced  in  the  smelting.  To  a 
charge  as  above  constituted  is  added  15%  or  more  of  coke,  not 
only  to  melt  the  charge  but  to  reduce  the  lead  oxide  to  metal  and 
the  ferric  oxide  to  the  ferrous  form. 


OF    THE    COMMON    METALS.  371 

123.     GENERAL     ARRANGEMENT     OF     SMELTING    WORKS. 

Fig.  155  is  the  plan  of  a  large  and  complete  silver-lead  works, 
the  Globe  plant  of  the  American  Smelting  &  Refining  Co.,  near 
Denver,  Colorado.  On  the  feed-floor  level  are  the  blast-furnaces,  the 
ore  beds,  the  sampling  building,  the  building  for  the  pot-roasting 
or  Huntington-Heberlein  process,  the  building  for  the  reverberatory- 
furnace,  and  the  sulphide-mill  where  the  sulphide  ores  are  crushed 
preparatory  to  roasting.  At  the  lower,  or  slag-floor,  level,  are  the 
boilers,  power-house,  and  settling-furnace  building.  The  buildings 
of  the  two  levels  form  a  complete  plant.  At  the  northeast  end 
of  the  works  is  situated  the  bag-house  where  the  flue-dust  is 
effectually  separated.  A  long  connecting  flue  leads  from  the  blast- 
furnace to  this.  Back  of  the  main  buildings,  and  now  not  altogether 
in  use,  are  the  Blake  and  the  Bruckner  roaster  buildings,  a  silver- 
lead  refinery,  and  a  parting-plant.  A  long  flue  leads  from  the 
reverberatory-roaster  building  and  from  the  H.-H.  building  to  the 
main  stack.  It  would  not  be  possible  to  turn  the  roaster  fume  into 
the  bag-house  because  of  the  sulphuric  acid  it  contains. 

124.     THE  SILVER-LEAD  BLAST-FURNACE. 

Fig.  156  represents  in  section  and  side  elevation,  a  water- 
jacketed  silver-lead  blast-furnace,  rectangular  in  plan,  44  in.  wide 
and  144  in.  long  inside  the  jackets  at  the  tuyere  level.  Fig.  157  is  a 
half  section,  and  end  elevation,  and  Fig.  158  is  a  view  of  the  same 
furnace.  The  furnace  comprises  a  crucible  standing  on  the  ground 
at  the  slag-floor  level,  the  shaft,  extending  from  the  crucible  to  the 
feed-floor,  and  the  stack  or  closed  top,  from  the  feed-floor  through 
the  roof  of  the  furnace  building. 

The  furnace  foundation  is  made  of  rubble  masonry,  or  of 
concrete.  It  extends  from  a  solid  footing  to  the  slag-floor  level  in 
a  solid  mass  over  sufficient  area  to  take  in  the  corner  posts.  Where 
another  furnace  is  in  operation  close  by,  the  slag  from  it  may,  with 
little  expense,  be  used  for  making  a  solid  slag-block  formation  by 
filling  the  foundation-pit  that  is  excavated  for  the  purpose. 

Upon  the  foundation  rests  the  crucible  shown  in  section  in  Fig. 
156  and  157  and  in  perspective  in  Fig.  158,  but  in  the  latter  the 
crucible  is  bound  with  steel-plates.  The  bottom  of  the  crucible  is 
a  steel  plate  8  by  16  ft.  in  area  on  which  rests  the  inclosing 
crucible-plates  of  cast-iron  securely  bound  with  rods.  Within  is 
built  the  fire-brick  crucible,  44  in.  wide  by  144  in.  long  inside  by 
2  ft.  deep,  the  sides  narrowing  downward.  At  one  side  is  a  channel 


372 


THE    METALLURGY    OF 


8  in.  square,  extending  from  the  sole  of  the  crucible  to  the  top 
level,  and  called  the  lead-well.     In  operation,  the  crucible  and  the 


Fig.    156.      SILVER-LEAD   BLAST-FURNACE    (LONGITUDINAL   ELEVATION). 

lead-well   are   full   of   molten   lead.     As   lead   accumulates   in   the 
crucible  it  rises  in  the  lead-well,  and  can  be  removed.    The  circular 


OF    THE    COMMON    METALS. 


373 


top-opening  of  the  lead-well  is  shown  at  the  front  in  Fig.  158.     The 
jackets  form  the  lower  part  or  bosh  of  the  shaft,  and  are  made  either 


Fig.    157.      SILVER-LEAD   BLAST-FURNACE    (TRANSVERSE   ELEVATION). 

of  cast-iron  or  of  steel.    As  seen  in  Fig.  156,  157,  and  158,  they  are 
6  ft.  high.    The  side-jackets  have  tuyeres,  making  a  bend,  or  knee, 


Fig.    158.      PERSPECTIVE   VIEW  OF   SILVER-LEAD  BLAST-FURNACE. 


OF    THE    COMMON    METALS. 


375 


in  the  jacket  as  shown  in  the  illustration  of  the  steel  side-jacket 
Pig.  159,  and  in  Fig.  157.  At  the  top  of  the  jacket  is  seen  the  spout 
into  which  the  pipe  enters  that  supplies  the  water,  and  at  the  front 


Fig.    159.      END  WATER-JACKET,   WITH   KNEE-BOSH. 


Fig.    160.      END   WATER-JACKET,    STRAIGHT. 

of  the  spout  is  seen  the  hole  for  the  outlet  water-pipe.  Thus  a 
circulation  of  water  through  the  jacket  is  insured.  Fig.  160  is  a 
steel  end-jacket  without  a  bosh.  The  jacket  is  shown  also  in  the 


376 


THE    METALLURGY 


longitudinal  view,  Fig.  156.  The  end-jackets  are  made  at  least  12  in. 
shorter  than  the  side-jackets  so  as  to  leave  room  below  for  the 
breast-opening.  This  space  is  generally  filled  with  a  tap- jacket. 
Jackets  are  also  made  of  cast-iron,  but  are  narrower  than  when  of 
steel.  A  side-jacket,  for  example,  taking  in  a  single  tuyere,  is  but 
18  to  20  in.  wide  (See  Fig.  161).  Each  jacket  has  a  spout  for 
receiving  and  discharging  water  as  is  the  case  with  steel  jackets. 
Hand-holes  near  the  bottom  are  put  in,  so  that  accumulated  scale, 
deposited  from  the  jacket-water,  can  be  removed.  Fig.  162  is  a  left- 


p 


u 
in. 


u 


Fig.   161.      CAST-IRON 
SIDE-JACKET. 


Fig.    162.      CAST-IRON 
END-JACKET. 


hand  end-jacket,  the  right-hand  one,  with  it,  forms  the  end  of  the 
furnace.  In  Fig.  156  is  seen  no  bosh  in  the  end-jacket.  Often, 
as  is  indicated  in  Fig.  15f,  a  bosh  is  provided.  Fig.  162  shows  a 
jacket  having  a  bosh,  and  indicates  the  method  by  which  it  is 
connected  with  the  side-bosh  by  a  rounded  corner.  The  object  of 
the  boshes  is  two-fold.  They  serve  to  support  the  charge,  and  they 
make  the  furnace  larger  above,  so  that  the  content  descends  slowly, 
and  remains  a  long  time  in  contact  with  the  reducing  gas.  Also 
when  accretions  form  in  the  shaft  there  is  room  thus  left  for  the 
smelting  operation. 

Air  is  supplied  to  the  tuyeres  from  the  bustle-pipe   shown  in 
Fig.  156  and  157.,  Canvas  sleeves  are  shown  by  which  the  connection 


OF    THE    COMMON"    METALS.  377 

is  made  between  the  bustle-pipe  and  the  tuyeres.  The  canvas  makes 
a  flexible  connection,  so  that  the  tuyeres  can  be  readily  removed, 
as  when  the  furnace  has  to  be  stopped  for  a  short  time.  In  fact, 
if  the  tuyeres  were  not  removed,  the  canvas  would  be  burned  by  the 
flame  from  the  furnace. 

A  system  of  water  pipes  supplies  the  jackets,  and  another  system 
takes  the  waste  water  from  them  to  funnels  at  each  corner  post 
(See  Fig.  158).  The  furnace  shaft,  extending  from  the  top  of  the 
jackets  to  the  feed-door,  is  lined  with  fire-brick,  the  whole  being 
firmly  tied  with  rods,  and  having  angle-irons  at  the  corners.  The 
brick  structure  of  the  furnace  is  supported  by  a  deck-plate  or 
mantel  resting  on  the  cast-iron  corner  posts  or  columns. 

Above  the  feed-floor  is  the  closed  top  or  stack,  which  collects 
the  smoke  and  gas  arising  from  the  charge,  and  delivers  it  to  the 
down-take,  a  5-ft.  pipe  that  rises  from  the  furnace-top  at  an  angle 
of  45°  then  descends  at  the  same  angle  to  the  main  flue.  Thus  no 
flue-dust  can  lodge  in  the  pipe ;  it  either  slides  back  into  the 
furnace  or  forward  into  the  flue.  When  the  furnace  is  run  down 
the  gas  becomes  too  hot  to  send  into  the  flue,  and  the  damper  at  the 
top  of  the  stack  is  opened  to  permit  the  escape.  At  the  charge- 
floor  level  are  large  doors  giving  access  to  the  furnace,  for  feeding 
and  cleaning  it  when  stopped  for  the  purpose. 

125.     BLOWING-IN  THE   BLAST-FURNACE. 

This  operation  may  be  divided  into  four  parts,  which  are : 
warming  the  crucible,  melting-in  the  lead,  filling  the  furnace,  and 
starting  the  smelting. 

Warming  the  crucible. — This  is  done  gradually.  The  breast  of 
the  furnace  being  open,  a  small  fire  with  light  wood  is  started  in 
the  crucible  to  expel  the  moisture  from  the  brickwork.  This  takes 
24  hours.  The  fire  should  be  regular,  the  wood  being  placed  at  the 
walls  of  the  crucible,  and  the  middle  left  clear.  When  the  fire  has 
burned  several  hours,  the  charcoal  and  ash  (both  non-conductors  of 
heat)  accumulate.  These  are  removed  together  with  the  glowing 
coals,  with  a  long  iron-handled  shovel,  and  a  new  fire  is  built. 
When  the  brickwork  seems  dry,  the  fire  is  increased  to  warm  it.  To 
do  this  the  tuyere  sacks  are  tied  up  so  that  no  air  passes  through 
them,  and  at  the  last  tuyere  a  length  of  tuyere-sacking  is  attached 
to  supply  air  to  a  6-ft.  piece  of  2%-in.  pipe.  The  blower  being 
set  in  slow  motion,  supplies  air  vigorously  through  the  blow-pipe 
to  consume  the  wood.  The  lead-well,  having  a  draft  toward  the 
crucible,  is  warmed  at  the  same  time  with  a  fire  of  wood  or  charcoal. 


378  THE    METALLURGY 

Particular  attention  is  given  to  warming  the  lead-well,  and  the 
heating  is  kept  up  until  the  outer  brickwork  of  the  crucible  feels 
warm  to  the  hand. 

Melting-in  the  lead. — The  hot  crucible  is  cleaned  out  again  and 
wood  is  put  in  which  soon  takes  fire.  Upon  the  wood  is  charged, 
with  a  paddle  or  'peel,'  a  dozen  bars  of  lead  (base-bullion  reserved 
from  the  last  run  of  the  furnace)  which  are  speedily  melted  by  using 
the  temporary  blow-pipe.  More  wood  is  added,  followed  by  other 
bars,  and  so  progressively  wood  is  burned  and  lead  is  melted.  When 
the  crucible  is  half  full,  the  fuel,  charcoal,  and  ash  are  removed,  a 
fresh  fire  is  made,  and  the  melting  is  completed.  It  takes  30,000 
Ib.  of  lead  to  fill  a  44  by  144-in.  furnace.  The  melting  begun  in 
the  evening  is  completed,  and  the  lead  skimmed,  when  the  day 
shift  comes  on  at  6  a.  m. 

Filling  the  furnace. — Dry  wood,  preferably  in  long  sticks  and 
split  4  in.  thick,  is  put  on  the  lead  bath.  As  much  is  used  as  can 
be  inserted  at  the  breast.  The  tap- jacket  is  next  set  in,  and  the 
breast  is  bricked  up,  while  more  wood  is  dropped  from  the  charge- 
door  to  fill  the  space  within  the  jacket  to  a  level  slightly  above 
the  tuyeres.  A  layer  of  charcoal,  two  feet  deep,  is  added,  followed 
by  coke  to  a  depth  of  12  to  18  in.  A  number  of  charges  of  fusible 
slag  follow,  each  with  the  regular  quantity  of  coke,  and  a  little 
iron  ore  and  limestone  to  flux  the  ash  of  the  coke.  Ore  charges 
next  replace  slag  charges  until  charging  is  normal.  Care  is  taken 
to  fill  the  furnace  evenly  and  quickly,  needing  the  personal  attention 
of  the  metallurgist  or  foreman  with  extra  men. 

Starting  the  smelting. — The  furnace  being  filled  half  up  to  the 
charge  floor  and  containing  a  few  ore  charges,  fire  is  started  at 
several  of  the  tuyeres  on  each  side  by  putting  in  a  little  greasy 
waste  and  lighting  it.  The  tuyeres  are  at  once  put  in  place  and  the 
blast  admitted  at  a  pressure  of  1  to  2  oz.  per  sq.  in.  The  wood  soon 
ignites  and  smoke  rises  through  the  charge.  The  blast  is  increased 
gradually  during  one  to  two  hours,  until  the  furnace  is  in  full 
operation.  As  the  slag  accumulates  it  is  tapped,  while  the  lead, 
accumulating  in  the  crucible  and  lead-well,  is  either  dipped  from 
the  lead-well  with  ladles,  or  tapped  through  a  lead  tap-hole  at  the 
level  of  the  top  of  the  crucible.  The  amount  of  lead  removed  at  one 
time  is  limited  to  1000  to  1500  Ib.  to  keep  the  crucible  always  full. 
The  operation  of  filling  and  starting  takes  7  hours. 

Regular  work  on  the  charge-flood. — This  consists  in  wheeling 
ore  and  fuel  from  the  bins  to  the  charge-scales,  weighing  the 


OF    THE    COMMON    METALS.  379 

required  amounts  into  charge-barrows  or  charge-cars,  then  dumping 
the  charge  thus  prepared  upon  the  feed-plates  in  front  of  the 
furnace-doors.  Every  material  except  the  foul  slag  is  weighed, 
and  even  the  slag  is  added  by  shoveling  in  regular  amounts.  A 
charge  is  dumped  on  one  charge-plate  and  the  coke  for  it  on  the 
other.  The  coke  is  fed  in  an  even  layer,  and  the  plate  thus  cleared 
receives  an  ore-charge.  The  ore-charge  on  the  other  charge-plate 
is  then  added,  taking  care  to  place  the  larger  ore  at  the  middle  of 
the  furnace  and  the  fine  at  the  walls  and  corners.  Such  a  distri- 
bution should  be  made  as  to  cause  the  smoke  and  gas  to  rise  evenly 
throughout.  The  blast  tends  to  ascend  at  the  walls  more  than  in 
the  middle,  but  by  the  distribution  it  is  compelled  to  rise  evenly. 
The  plate,  now  cleared  of  the  ore-charge,  receives  the  coke  that  is 
next  to  go  in.  The  content  of  the  furnace  remains  at  the  level  of  the 
charge-floor,  new  charges  being  supplied  as  the  surface  sinks. 

Regular  work  at  the  ground  or  slag-floor. — This  consists  in  regu- 
lating the  water-supply  at  the  jackets,  seeing  that  the  tuyeres  are 
clean  and  open,  tapping  and  stopping  the  slag,  and  when  the  slog 
and  matte  are  separated  in  a  fore-hearth,  tapping  the  matte,  placing 
the  slag-pots  (See  Fig.  135),  and  removing  them  when  full  to  the 
dump,  tapping  or  dipping  the  lead  (base-bullion)  into  molds,  and 
removing  and  piling  up  the  bars  of  lead  that  later  are  to  be  sampled. 

The  matte  is  tapped  into  pots  or  molds,  and  when  solidified,  is 
broken  and  sent  to  be  crushed  to  3-mesh  size  for  roasting.  The  slag, 
as  in  copper  practice,  may  be  granulated  and  removed  in  a  launder 
by  means  of  running  water.  There  is  the  objection  to  this,  however, 
that,  by  reason  of  irregular  working,  matte  may  be  carried  away  and 
never  recovered  from  slag  once  granulated;  hence  removal  in  slag- 
pots  is  preferred. 

When  a  pot  of  slag  has  been  taken  to  the  edge  of  the  dump  and 
has  stood  there  a  few  minutes,  a  shell  or  *  skull '  of  solidified  slag  half- 
an-inch  thick  forms.  A  hole  is  broken  through  the  top  and  the  molten 
content  is  poured  out.  The  shell  is  returned  to  the  furnace  for  re- 
smelting.  It  is  found  to  contain  drops  of  matte  that  have  settled 
from  the  molten  slag  and  these  are  recovered  only  by  smelting  it. 
Slag  is  also  conveniently  removed  by  means  of  large  slag-pots 
mounted  on  trucks  and  moved  about  by  a  horse  or  by  an  industrial 
locomotive,  Fig.  137.  At  certain  large  silver-lead  smelting-works 
the  slag  and  matte  together  are  taken  to  a  reverberatory  melting- 
furnace,  called  a  settling-furnace,  or  settler,  where  the  matte 
thoroughly  settles  from  the  slag,  and  is  tapped  just  above  the  bottom 
of  thft  furnace.  The  bath  of  slag  is  maintained  18  to  24  in.  deep  in 


380  THE    METALLURGY 

the  furnace,  and  when  accumulated,  is  tapped  at  a  level  a  little  be- 
low the  surface,  and  delivered  to  a  train  of  slag-cars  operated  by  a 
locomotive.  A  fire  with  a  smoky  reducing  flame  is  maintained  in  the 
settler.  This  gives  a  slag,  cleaner  by  0.3%  lead  and  0.3  oz.  silver  per 
ton  than  the  ordinary  method ;  but  it  needs  the  combined  output  of 
several  furnaces  to  keep  it  hot  and  in  good  working  condition. 
Where  much  zinc  is  in  the  charge  the  slag  also  may  give  trouble. 

126.     CHEMISTRY  OF  THE  BLAST-FURNACE. 

The  surface  of  the  charge  should  look  dead,  showing  no  visible 
heat  or  flame  (over-fire).  With  much  over-fire,  there  is  a  loss  of 
lead  due  to  volatilization.  The  moisture  in  the  charge  soon  dries  at 
the  temperature  of  the  rising  gases  (200°C.).  The  heat  thus  ab- 
sorbed is  small,  and  by  calculation  it  is  found  to  be  but  one-thirtieth 
of  the  total  supplied  by  the  fuel  when  5%  moisture  is  in  the  charge. 
As  the  materials  of  the  charge  descend  in  the  furnace,  carbon  di- 
oxide begins  to  be  driven  from  the  limestone,  and  the  iron  reduces 
from  the  ferric  to  the  ferrous  form.  Half  way  down  at  a  tempera- 
ture of  800°C.  the  reactions  are  complete.  The  lead  in  oxidized 
form  is  reduced  by  the  CO  of  the  gas  and  by  the  red-hot  coke.  Ga- 
lena reacts  with  the  iron  oxide  and  carbon  as  follows : 

(1)  Pl?+  FeO  +  C  ==  Pb  +  FeS  +  CO 

Sometimes  scrap-iron  is  added  to  the  charge,  and  acts  with  galena 
or  other  sulphides  as  in  the  nail  assay  for  silver  as  follows: 

(2)  PbS  +  Fe  =  FeS  +  Pb 

As  the  lead,  thus  reduced,  drops  through  the  charge,  it  collects 
the  gold  and  silver  as  well  as  a  part  of  the  arsenic  and  antimony  and 
enters  the  crucible  as  base-bullion.  When  antimony  is  present  it  is 
reduced  like  lead  and  alloys  with  the  latter.  Anglesite  is  reduced  by 
contact  with  the  fuel  and  iron  oxide  according  to  the  following  re- 
action : 

(3)  PbSO4  +  FeO  +  5C  =  Pb  +  FeS  +  5CO 

Where  much  angelsite  is  in  the  charge  more  than  the  usual  quantity 
of  fuel  is  demanded.  Since  the  affinity  of  sulphur  for  copper  is 
greater  than  for  iron,  copper  sulphide  remains  unreduced,  and  cop- 
per as  an  oxide  takes  sulphur  from  the  charge  and  with  the  iron 
forms  matte.  Lead  to  the  extent  of  10  to  20%,  either  as  sulphide  or 
in  metallic  form,  is  also  taken  up  by  the  matte.  To  some  extent  zinc 
sulphide  also  enters  the  matte.  The  ferrous  oxide,  not  needed  to 
satisfy  the  matte,  and  the  CaO  and  MgO  in  the  charge,  must  be  pres- 
ent in  sufficient  quantity  to  form  a  suitable  fusible  slag  with  the 


OF    THE    COMMON    METALS.  381 

silica.  Oxidized  arsenic  compounds  react  with  iron  and  carbon  pro- 
ducing a  speiss,  often  of  the  form  Fe4As,  and  requires  extra  fuel. 
The  reaction  is  as  follows  : 

(4)     As,03  +  8FeO  -f  11C  =  2Fe4As  +  11CO 

The  molten  products  separate  at  the  hearth  according  to  the  spe- 
cific gravity,  that  of  lead  being  11.5,  speiss  6.0,  matte  5.2,  and  slag 
3.6.  The  lead,  collected  in  the  crucible,  is  withdrawn  at  the  lead- 
well.  The  slag,  matte,  and  speiss  are  drawn  off  at  the  slag  tap-hole 
on  the  level  of  the  top  of  the  crucible  and  of  the  lead.  The  separa- 
tion of  matte  from  slag  is  generally  effected  in  a  fore-hearth  outside 
the  furnace. 

127.     SLAGS  IN  SILVER-LEAD  SMELTING. 

The  object  of  silver-lead  smelting  is  to  reduce  the  lead,  and  inci- 
dentally the  gold  and  silver,  from  the  ore.  The  sulphur  present 
forms,  with  the  copper,  iron,  and  a  part  of  the  lead,  a  complex  arti- 
ficial sulphide  (matte),  while  basic  flux  is  added  to  form  a  slag  of 
the  composition  that  experience  shows  necessary. 

Slags  are  silicates  of  extraordinary  complexity;  and  not  all 
merely  fusible  slags  work  well  in  a  silver-lead  blast-furnace.  Type 
slags  are  those  so  proportioned  in  silica,  iron  oxide,  and  lime,  as  to 
work  well  in  the  blast-furnace.  Slags  that  vary  from  the  proportions 
become  defective  in  operation.  To  fulfil  the  requirements  of  good 
slag,  it  should  have,  in  the  normal  operation  of  the  furnace,  not 
more  than  0.7%  lead,  or  0.5  oz.  silver  per  ton,  when  producing  base- 
bullion  not  higher  than  300  oz.  silver  per  ton.  The  density  should 
not  be  greater  than  3.6.  It  should  not  permit  accretions  to  form  at 
the  hearth,  nor  the  creeping-up  or  appearance  of  over-fire.  If  a  slag 
varies  from  one  of  the  types  given,  it  is  either  poorly  reduced  or 
makes  other  trouble  in  the  furnace.  Thus  a  slag  of  the  three-quarter 
type,  in  which  the  CaO  falls  to  20%  (the  other  constituents  as 
given  in  the  table),  is  found  to  contain,  for  example,.  1%  lead,  and 
more  than  1  oz.  silver  per  ton;  that  is;  it  is  'dirty.'  It  easily 
may  happen  that  a  dirty  slag  is  fusible,  but  we  know  that  the  slag 
will  work  satisfactorily  if  correct  according  to  the  type,  the  other 
conditions  of  good  running  being  in  evidence.  Indeed  it  is  a  common 
experience  that  a  furnace,  working  poorly  on  an  incorrect  slag, 
begins  to  run  well  when  a  correct  slag  comes  down.  Following 
we  give  a  table  of  type-slags  that  have  been  found  to  work  well 
in  practice. 


382  THE    METALLURGY 

TABLE   OF  TYPICAL   SLAGS. 

SiO,    Fe(Mn)O  Ca(Ba,Mg)O 

Type.  %  %  % 

Quarter-slag C  28  50  12 

Silicious  quarter-slag H  32  47  11 

Half-slag    E  30  40  20 

Half-slag    J  31  38  21 

Silicious    half-slag    I  35  38  17 

Three-quarter-slag    F  33  33  23 

Silicious   three-quarter-slag    M  36  31  23 

Whole,  or  1  to  1  slag G  35  27  28 

According  to  the  ratio  of  FeO  to  CaO  the  slag  is  called  a  '  quarter, ' 
a  'half,'  or  a  'one-to-one'  slag,  etc.  Thus  the  slag  E  of  the  sub- 
joined table  is  called  a  half-slag,  the  CaO  being  but  half  of  the 
FeO.  The  slag  C  is  a  quarter-slag,  the  CaO  being  quarter  of  the 
FeO.  In  this  table  of  typical  well-tested  slags  the  three  elements 
SiO2,  Fe(Mn)O,  and  Ca(Mg,Ba)0  are  calculated  to  comprise  90%. 
If  the  sum  varies  from  this,  the  ratio  is  still  to  be  preserved. 

Since  any  of  the  slags  of  the  table  can  be  used,  the  question 
arises,  which  is  to  be  chosen.  In  this  we  are  guided  by  the  economic 
conditions.  If  the  ore  of  the  district  is  silicious,  and  most  profit 
is  derived  by  treating  the  ore  at  hand,  use  a  silicious  slag  that 
needs  the  smaller  amount  of  flux.  If  irony  or  limey  ores  are  plentiful 
and  profitable  to  smelt  we  use  them,  substituting  them  for  flux. 
It  is  found,  however,  that  slags  of  the  type  M  and  G  of  the  table 
drive  more  slowly,  and  require  more  fuel  than  the  basic  ones. 
Perhaps  the  most  satisfactory  of  the  slags,  and  the  one  that  can 
be  used  where  silicious  ores  are  plentiful,  is  the  three-quarter  slag, 
F.  When  a  slag  of  a  certain  type,  for  example  a  silicious  one, 
is  not  working  well  and  is  forming  accretions  in  the  furnace,  a 
radical  change  to  a  basic  type  is  found  beneficial,  or  from  a  basic 
slag  to  a  silicious  one. 

128.     ACTION  OF  VARIOUS  BASES  IN  SLAGS. 

Iron. — Iron  ore  is  quickly  reduced  to  ferrous  form  under  the 
action  of  the  CO  in  the  furnace  or  of  the  highly  heated  fuel  thus : 

(1)  Fe203  +  C&;  2FeO  -f  C02 

Iron  oxide,  being  a  stronger  base  than  lead  oxide,  replaces  it  in 
the  slag,  and  the  latter  is  reduced  by  carbon  to  metallic  lead. 

(2)  PbSi03  +  FeO  +  C  =  FeSiO3  +  Pb  +  CO 
Manganese. — The  equivalent  for  manganese  is  55,  and  for  iron 

">6.    and   they   are   reckoned    as   having   equal   values   for   fluxing. 


OF    THE    COMMON   METALS.  383 

Manganese  is  found  in  some  of  the  Leadville  iron  ores  to  the  extent 
of  10  to  15%,  and  since  by  introducing  another  element,  it  adds 
to  the  complexity  of  the  slag,  it  also  adds  to  the  fusibility. 

The  alkaline  earths. — Lime,  magnesia,  and  baryta  act  in  inverse 
ratio  to  the  atomic  weights  in  fluxing  silica.  Hence  to  obtain  the 
equivalent  in  lime,  the  amount  of  lim^  needed,  c^prconcd  in-  per 
^  is  multiplied  by  1.4  for  magnooia,  and  by  0.4  for  baryta. 
A  slag,  high  in  lime  and  consequently  low  in  iron  like  the  last 
three  in  the  table,  is  of  low  specific  gravity.  Its  use  thus  results 
in  a  better  separation  of  slag  from  the  heavier  matte.  Lime  being 
a  stronger  base  by  one  half  than  iron,  and  generally  a  cheaper 
flux,  the  tendency  is  to  choose  the  limey  slags.  It  is  noticed  that 
the  higher  the  silica  content  of  the  slags  of  the  table  the  higher  is 
the  lime,  and  that  high  silica  calls  for  high  lime.  Dolomite,  having 
a  high  content  in  magnesia,  generally  is  avoided  in  silver-lead 
smelting,  for  it  tends  to  make  slag  pasty  and  streaky,  and  the 
unfavorable  effect  is  aggravated  when  zinc  also  is  present.  Two 
analyses  of  limestone  and  of  dolomite  are  given  below  to  show 
conditions  typical  of  actual  practice. 

Canyon  City  limestone:  CaO,  49.8%;  MgO,  3.0%;  SiO2,  3.1%; 
Fe,  0.8%. 

Iron  county,  Missouri,  dolomite:  CaO,  26.6%  ;  MgO,  17.6%  ;  SiOo, 
5.1%;  Fe,  3.3%. 

Flourspar. — This  has  no  unfavorable,  but  rather  a  favorable 
effect  upon  the  quality  of  the  slag.  The  fluorine,  however,  uses 
CaO,  and  hence  the  slag  must  analyse  higher  in  CaO  than  the  type 
requires,  or  it  will  not  be  clean. 

Alumina. — It  is  uncertain  whether  alumina  acts  as  an  acid  or 
a  base.  It  is  sufficient  for  the  purpose  of  silver-lead  smelting  to 
regard  it  as  a  neutral  constituent  that  dissolves  in  slag  and  acts 
in  neither  way. 

Zinc. — Either  blende  or  zinc  oxide  causes  difficulties  in  the 
blast-furnace,  the  blende  being  the  more  objectionable.  Blende  is 
in  part  decomposed  in  the  presence  of  iron  to  zinc  oxide,  but  the 
zinc  in  any  form  tends  to  make  a  stiff,  pasty,  difficultly  fusible 
slag.  It  may  be  regarded,  like  alumina,  as  being  dissolved  in  the 
slag.  It  goes  into  both  the  slag  and  the  matte,  and  diminishing 
the  specific  gravity  of  the  latter  it  causes  a  less  perfect  separation 
of  the  two.  Where  much  zinc  is  in  the  charge,  it  is  customary  to 
modify  the  type-slag  by  calculating  the  zinc  oxide  as  replacing  one 
half  the  percentage  of  lime.  Take  for  example  the  half-slag  J 
of  the  table. 


384  THE    METALLURGY 

Without         With     Recalculated 


Zinc 

Zinc 

Zinc 

Per  cent. 

Per  cent. 

Per  cent. 

SiO,     

31 

31 

•     29.5 

FeO 

38 

38 

36.0 

CaO     . 

21 

17 

16.0 

ZnO 

8 

7.5 

90  94  90.0 

In  the  first  column  we  write  the  slag  as  the  type  requires.  In  the 
second  column  we  add  the  8%  Zn  and  reduce  the  lime  by  4%  by 
which  the  total  becomes  94.  Since  the  constituents  should  amount 
to  but  90%,  all  are  reduced  proportionally  in  the  third  column 
so  as  to  give  90%  as  the  sum. 

Copper. — Copper  present  in  the  charge  enters  the  matte  when, 
as  generally  is  the  case,  sulphur  is  present  with  which  it  can  com- 
bine. In  smelting  carbonate  or  oxidized  ores,  which  furnish  no 
sulphur,  the  copper  becomes  reduced,  and  enters  the  base-bullion, 
giving  a  lead  so  drossy  sometimes  as  to  clog  the  lead-well,  and 
accumulate  and  solidify  in  the  crucible.  The  remedy  is  to  supply 
sulphide  to  form  matte  into  which  the  copper  can  enter. 

Antimony. — Either  as  an  oxide  or  a  sulphide,  antimony  is 
reduced  like  lead.  It  alloys  with  the  base-bullion,  making  it  hard, 
and  is  removed  and  recovered  later  in  refining  the  base-bullion. 

Arsenic. — This  frequently  is  encountered  in  silver-lead  smelting. 
When  present  in  small  quantity  it  is  volatilized,  but  in  large 
quantity  it  forms  a  speiss.  Where  it  is  intended  to  produce  a  speiss, 
iron  is  provided  with  which  the  arsenic  unites.  In  the  fire-assay 
of  arsenic-bearing  lead  ores,  a  bead  of  speiss  is  found  attached  to 
the  lead  button.  From  the  percentage  of  this  we  can  compute  the 
weight  of  the  speiss  that  will  be  formed:  and  we  may  assume  that 
70%  of  it  is  Fe.  Where  a  direct  determination  of  arsenic  is  made 
we  can  compute  the  weight,  and  multiply  this  by  2.3  to  express 
the  quantity  of  Fe  to  be  provided  on  the  charge  for  the  purpose. 

129.     FUEL  IN  SILVER-LEAD  SMELTING. 

The  fuels  used  in  silver-lead  smelting  are  coke,  charcoal,  or  a 
mixture  of  the  two.  Wood  and  hard  coal  have  been  used  experi- 
mentally, the  former  in  certain  cases  of  scarcity  of  fuel. 

Coke. — Coke  is  the  kind  of  fuel  commonly  used.  The  ash  varies 
from  10  to  22%,  and  the  fixed  carbon  from  89  to  77%.  In  coke  of 


OF   THE    COMMON"   METALS.  385 

high  ash,  not  only  is  the  ash  to  be  smelted,  but  the  carbon  is 
correspondingly  low,  so  that  the  coke  is  less  efficient.  A  great 
difficulty  with  high-ash  coke  is  that  it  is  often  friable,  making 
accretions  or  scaffolds.  Analyses  of  two  typical  samples  of  coke 
give  the  following  results  : 

Connellsville  coke  contains  fixed  carbon  87.5%,  ash  11.3%,  and 
sulphur  0.7%.  El  Moro  coke,  fixed  carbon  77.0%,  ash  22.0%,  and 
sulphur  (when  the  coke  is  made  from  unwashed  coal)  0.9  per  cent. 

In  computing  a  charge  the  coke-ash  is  taken  into  account, 
analyses  being  as  follows :  Ash  of  Connellsville  coke :  Si02,  44.6%  ; 
Fe,  15.9%;  CaO,  7.0%;  MgO,  1.9%;  ash  of  El  Moro  coke:  Si02, 
84.5%,  and  Fe,  5.0%. 

Charcoal. — This  fuel  is  used  in  districts  far  from  railroads,  where 
the  cost  of  coke  is  high.  It  is  a  good  fuel  for  oxidized  ores,  but 
is  friable  and  makes  undesirable  fine  which  may  form  accretions 
or  scaffolds  in  the  furnace.  It  renders  a  charge  more  open  than 
coke,  and  contains  less  than  2%  ash.  Coke  weighs  25  Ib.  and 
charcoal  10  Ib.  per  cu.  ft.  when  loose,  the  weight  of  a  bushel  of 
charcoal  being  14  to  16  Ib.  Even  where  charcoal  is  cheap  it  is 
desirable  in  operating  the  furnace  to  use  part  coke  which,  fed  to 
the  walls,  burns  more  slowly  than  charcoal  and  makes  the  tuyere- 
zone  hotter  and  gives  a  more  liquid  slag. 

Quantity  of  fuel. — This  varies  according  to  the  nature  of  the 
charge,  and  generally  is  from  12  to  15%.  Charges  that  contain 
sulphur  and  make  matte  need  less  fuel  than  oxidized  ores.  Only 
sufficient  is  used  to  give  adequate  reduction  and  a  hot  slag;  and 
the  metallurgist  is  guided  by  these  requirements  in  adding  the  fuel. 

130.     CALCULATION  OF  A  CHARGE. 

When  sulphur-bearing,  oxidized  or  silicious  ore  is  used,  we 
have  to  consider  not  only  the  sulphur,  silica,  and  other  constituents 
of  the  ore,  but  also  the  products  of  the  furnace  that  remove,  the 
constituents. 

Ore  (galena  for  example)  containing  less  than  10  to  12%  sulphur 
generally  is  smelted  without  roasting.  It  is  cheaper  to  do  this, 
for  by  roasting,  the  sulphur  of  the  ore  is  reduced  to  but  3  to  4%. 
Many  ores  within  the  above  limit  are  leady  ores,  and  difficult  to 
roast  because  of  the  fusible  nature,  but  the  matte  that  they  produce 
is  easy  to  roast  for  the  elimination  of  sulphur. 

Ore  intended  for  roasting  may  be  simple,  consisting  of  iron 
sulphide,  or  complex  as  shown  by  the  following  analysis  of  a  roasted 
ore;  Si02,  10 %  •  Fe  and  Mn,  27%;  CaO,  MgQ,  and  BaO,  2%; 


386  THE    METALLURGY 

Zn,  8.8%  ;  Cu,  0.4%,  S,  6%  ;  Pb,  35%,  and  Ag,  50  oz.  per  ton.     The 
base  in  the  roasted  ore  was  present  as  sulphide  in  the  raw  ore. 

The  so-called  oxidized  ores  consist  of  the  carbonate  of  lead 
with  a  gangue  of  iron  oxide,  limestone,  dolomite,  and  silica.  Such 
ores  though  called  oxidized,  often  contain  a  little  sulphur,  as 
sulphide  (galena  or  pyrite)  or  as  sulphate. 

Silicious  ores  are  added  to  charges,  in  spite  of  the  large  excess 
of  silica,  because  the  gold  and  silver  are  present  in  quantity  to 
pay  to  recover.  The  lead  of  the  charge  takes  the  gold  and  silver 
contained  in  such  ore,  while  the  silicious  gangue  is  fluxed  into  a 
barren  slag  and  sent  to  waste. 

Both  iron  ore  and  limestone  are  added  to  the  charge  for  fluxing 
the  silica,  making  a  slag  of  a  predetermined  composition  or  type. 
If  the  fluxes  contain  gold  or  silver  the  metals  can  be  recovered, 
since  they  go  into  the  base-bullion  or  work-lead.  Without  gold  or 
silver  they  are  called  barren  or  'dead'  fluxes.  Ore  carrying  an 
excess  of  iron  or  lime  over  silica  (called  iron  or  lime  excess)  is  in 
the  same  category,  since  the  excess  is  useful  for  fluxing,  and  is 
credited  in  purchasing  ores.  Thus,  ore  containing  10%  SiO2  and 
40%  Fe  is  said  to  carry  30%  iron  excess.  , 

Not  all  the  slag  that  issues  from  the  furnace  is  clean.  At  the 
spout  where  the  matte  flows,  and  in  the  shell  lining  the  cavity  of 
the  fore-hearth,  slag,  containing  drops  of  lead  and  matte,  is  found. 
When  a  slag-pot  is  emptied  at  the  edge  of  the  dump,  there  remains 
a  shell  or  coating  of  solidified  slag.  This  shell,  half  an  inch  thick, 
is  found  to  contain  drops  of  matte  that  did  not  entirely  settle  in 
the  fore-hearth.  This  is  particularly  true  when  the  fore-hearth  has 
formed  a  thick  lining  and  soon  must  be  replaced  by  another.  All 
this  slag  having  value,  and  called  'foul  slag,'  is  an  acceptable 
addition  to  the  charge  because  of  the  fusibility,  and  the  coarse 
condition,  permitting  free  passage  of  the  air  of  the  blast. 

Computation  of  the  charge. — To  determine  the  amount  of  the 
fluxes  (iron  ore  and  limestone)  to  add  to  the  charge,  to  give  a 
slag  of  a  desired  composition,  it  is  necessary  to  know  the  weight 
of  the  ores  to  be  used,  and  the  results  of  analysis  of  the  ore,  fluxes 
and  fuel,  and  also  the  composition  of  the  slag  and  matte  that  are 
to  be  produced. 

First  example. — Following  is  an  example  of  a  charge-sheet  for 
the  calculation  of  a  charge  containing  no  roasted  ore.  The  charge 
is  to  weight  approximately  1000  lb.?  a  quantity  sufficient  to  fill  a 
charge-buggy. 


OF    THE    COMMON    METALS. 


387 


CHARGE   SHEET. 


Name  of  Ore 

Weigut 

Pb 

SiO2 

Feaud 
Mn 

CaO  and 
MgO 

S 

^g 

H20 

Wet 

Dry 

% 

Wt 

%    Wt 

% 

Wt.     % 

Wt. 

% 

Wt 

Oz 

Chrysolite  No.  28 

3.0 

515 

500 

21.0 

105 

32.0    160 

15.0 

75 

100 

50 

4.4 

22.0 

75 

18.7 

Ontario  
Iron  Ore 

5.0 
2.0 

52 

20  1 
200 
150 

50 
200 
200 

105 

60.0     30 
4.0       1 
30       6 
4.0       6 
203 

20 
57.0 
3.5 

2.6 

1 
114 

7 
4 

2.0 
3.0 
54.0 
1.5 

1 

6 
108 
2_ 
167 

0.5 
0.7 

1.1 
23.1 

75 

1.9 
20.6 

Limestone 

Coke  

950 

201 

Fe  for  Matte  = 

_S_j 

U  OS  for  Slag.  etc. 

Fe  for  Slag 

168 

12.0  S  for  Matte 

V2    Slag    SiO,  =  30.0  =  1.00   factor 

FeO  =  40.0  =  31.1   Fe  =  1.04 
CaO  =  20.0  =  0.67 


Matte    S  =  20.0 
Fe  =  55.0 
Cu  =     5.0 
Pb  =  15.0 
90.0  Fe*        55 

1  —  —  —  2.75   factor 

S  20 

The  items  and  the  analyses  of  the  charge  are  written  in.  the 
proper  columns.  Lead  ore,  'Chrysolite,  lot  No.  28'  is  used,  to  give 
approximately  10%  or  100  lb.  of  lead.  To  this  is  added  silicious 
'Ontario  ore'  with  the  fluxes,  to  make  up  the  remaining  500  lb. 
as  experience  shows  necessary.  If  upon  calculation  we  find  that 
too  little  silicious  ore,  and  hence  too  little  flux  has  been  taken  to 
make  up  this  500  lb.,  we  can  increase  it  as  needed. 
t  Of  the  sulphur,  we  assume  that  20%  is  volatilized,  and  that  the 
yslag  retains  fffo  ^s  weight  of  sulphur.  From  the  silica  we  compute 
the  slag  to  be  equal  to  ||^  or  6gD  lb.  Neglecting  fractions  we 
then  have  : 

0.  >-J»7£  Pounds. 

Sulphur  in  slag>$f  of  600  lb 6 

volatilized  20%  of  23.1  lb 5 

remaining  for  matte 12 


23 

Multiplying  the  12  lb.  sulphur  in  the  matte  by  the  factor  2.75, 
which  expresses  the  ratio  of  sulphur  to  iron,  we  find  33  lb.  to  be 
the  amount  of  iron  needed  for  matte.  This  leaves  168  lb.  iron  for 
the  slag. 

We  have  chosen  a  half-slag  for  this  particular  case,  and  find 
from  it  a  factor  of  (  ?i  =  1.04)  expressing  the  ratio  of  silica  to 


388  THE    METALLURGY 

-.. 

iron.  Multiplying  the  weight  of  the  silica  by  this  factor  gives  211 
Ib.  iron  needed.  We  have,  however,  but  171  lb.,  and  the  difference 
can  be  made  up  by  increasing  the  iron  by  40  lb.,  or  the  iron  ore  by 
80  lb.,  since  the  iron  ore  is  approximately  one-half  iron.  Again, 
multiplying  the  factor  expressing  the  ratio  of  silica  to  lime  (0.67) 
by  the  weight  of  silica,  we  find  136  lb.  needed,  so  that  we  have  31 
lb.  too  much.  Now,  since  the  limestone  is  approximately  one-half 
CaO,  we  diminish  the  limestone  60  lb.  Erasing  the  old  values  and 
substituting  the  new  ones  we  carry  the  calculation  through  again. 
The  results  should  be  correct  within  a  few  pounds,  and  may  be 
corrected  again  if  not  so.  Fractions  of  pounds  are  neglected,  and 
the  weight  of  the  fluxes  need  not  be  written  nearer  than  10  pounds, 
since  variation  in  the  charge,  due  to  variation  in  the  ore  and 
weighing,  easily  may  be  greater  than  fractions  of  10  pounds. 

The  slag  actually  produced  may  vary  from  the  desired  composi- 
tion. When  a  new  charge  has  been  put  on,  and  after  several  hours 
the  slag  has  come  down,  that  is,  has  begun  to  flow  from  the  furnace, 
it  is  analyzed  and  the  result  known  in  two  or  three  hours.  The 
charge  is  then  altered  to  give  a  correct  slag.  Before  making  any 
changes  we  must  be  sure  that  the  slag  is  hot  and  well  reduced. 
The  total  weight  of  the  charge  can  be  varied  conveniently  by  varying 
the  amount  of  silicious  ore,  retaining  unaltered  the  weight  of  the 
lead. 

Second  example. — For  a  charge  that  is  to  contain  roasted  ore  or 
other  sulphur-bearing  material,  the  following  sheet  gives  a  good 
illustrative  example. 

We  assume  that  15%  of  fuel  is  to  be  used,  and  that  the  weight 
of  the  charge  is  to  be  1000  lb.  Let  us  say  that  we  have  ores  coming 
in  proportions  to  permit  using  them  in  the  ratio  indicated  by  the 
first  two  items  of  the  charge-sheet  (300  lb.  roasted  ore  and  200  lb. 
lead  ore).  We  can  make  up  the  remaining  500  lb.  with  silicious 
ore  and  flux,  in  the  proportion  that  experience  suggests,  being 
assured  that  the  lead-bearing  ore  furnishes  about  100  lb.  lead,  or 
10%  of  the  charge.  We  find,  including  the  lead  in  the  silicious 
ore,  that  we  have  104  lb.  total  lead  which  meets  the  requirement. 
Next  enter  the  150  lb.  coke,  estimating  the  percentage  composi- 
tion from  the  following  data  (for  example  of  Connellsville  coke) 
with  11%  ash  of  Si02,  44%;  Fe203,  22.7%;  CaO,  7.3%;  MgO, 
1.9%.  Since  the  iron  has  been  reported  in  ferric  form,  we  have 
to  multiply  the  22.7%  Fe2O3  by  0.7  to  obtain  the  equivalent  iron. 
Magnesia  being  the  stronger  base,  we  multiply  the  percentage 
(1.9%)  by  1.4  to  obtain  the  lime-equivalent  of  the  magnesia,  adding 


OF    THE    COMMON    METALS. 


389 


it  then  to  the  lime,  making  in  all  11%.  Estimating  this  for  the 
coke,  by  the  percent  of  the  percent,  we  have  Si02,  4.9%  ;  Fe,  1.7% ; 
Ca  Oand  MgO  1.1  per  cent. 


CHARGE   SHEET. 


Name  of  Ore 

Weight 

Pb 

S1O2 

Fe  and 
Mn 

CaO  and 
MgO 

S 

H2O 

Wet 

Dry 

% 

Wt 

% 

Wt 

% 

Wt. 

% 

Wt. 

% 

Wt 

Roasted  Ore  
Lead  Ore. 

50 

300 
200 
200 
100 
200 

1000 

12.0 
25.8 
8.0 

36 
52 
16 

104 

12.0 
22.6 
63.3 
5.0 
3.5 
4.9 

48 
45 
127 
10 
3 
7 
240 

32.0 
18.1 
16.0 
540 

1.7 

96 
36 
32 
51 

3_ 
221 

2.0 
5.4 

52.0 
1.1 

6 
11 

101 
2_ 
123 

5.0 

2.0 
30 

0.7 

15 
4 
6 

1 
26 

Silicious  Ore  
Iron  Ore 

Limestone  
Coke 

Fefor 
Matte      - 

=   37 

S  volatil- 
ized 

5.0 

FeforSlag=l£4 

S  in  Slag     6.0 

11 

15 

For 
Matte 

Slag. 

SiOo  =33% 

FeO  =33% 

CaO  =  24% 

Other   bases  =  10% 


Fe  =  25.7  =  0.77  X  SiO2 
=  0.73  X  Si02 


Matte. 

S  =  20% 

20%  S  X  2.5  =  Fe  =  50% 

Cu  — .    5% 

Pb  =  157o 


100% 

Slag  =  730  Ib. 
S  in  slag  =  (730  X  0.8%)  =  6  Ib. 

Write  in  the  constituents  in  percentage  and  pounds,  footing  up 
the  columns  to  obtain  the  total  for  each  constituent.  All  fractions 
are  neglected.  At  the  lower  left  corner  of  the  charge-sheet  write 
the  type  of  slag  chosen  from  the  list  of  type-slags,  section  127.  The 
type  chosen  depends  upon  the  kind  found  commercially  to  be  the 
most  profitable.  This  again  depends  upon  the  cost  of  flux  and 
upon  the  ores  that  give  the  most  profit  in  treatment.  Where  it 
pays  best  to  treat  silicious  ore,  one  uses  a  silicious  slag,  but  where 
there  is  more  profit  in  an  irony  ore,  then  a  basic  slag  is  preferred. 
For  the  present  calculation  we  choose  the  three-quarter  slag,  F, 
containing  Si02,  33%  ;  FeO  and  MnO,  33%,  and  CaO  and  MgO, 
24%.  The  other  bases  such  as  ZnO,  A12O3,  alkalis  and  sulphur 
make  up  the  remaining  10  per  cent. 

The  33%  FeO  is  calculated  to  Fe,  making  25.7%.  Finding  the 
ratio  of  Si02  to  Fe,  we  get  the  factor  0.77.  For  CaO  we  obtain,  in 
the  same  way,  the  factor  0.73.  On  the  other  side  of  the  sheet 


390  THE    METALLURGY 

tabulate  a  matte  analysis,  and  calculate  the  factor  expressing  the 
ratio  of  S  to  Fe,  or  2.5. 

First  consider  the  sulphur.  Experience  shows  that,  of  the 
amount  of  sulphur  present,  we  can  estimate  that  20%  or  5  Ib.  will 
be  volatilized,  and  also  that  the  slag  will  contain  0.8%  S.  The 
weight  of  the  slag  is  found  by  dividing  the  weight  (240  Ib.)  by  the 
percentage  expressed  decimally  (0.33).  This  equals  730  Ib.  Hence 
there  remains  15  Ib.  of  sulphur  to  enter  the  matte,  and  this  multiplied 
by  the  factor  2.5  gives  37  Ib.  Fe.  This  subtracted  from  the  total  Fe 
leaves  183  Ib.  for  the  slag.  Multiplying  the  240  Ib.  of  silica  by  the 
factor  0.77,  we  find  185  Ib.  Fe,  so  that  we  are  close  to  the  calculated 
iron  needed.  Again,  calculating  the  lime  needed  (240  multiplied  by 
0.73  or  175  Ib.)  we  find  we  are  short  (175  minus  123)  or  52  Ib.,  equal 
to  100  Ib.  of  limestone  of  50%  CaO.  This  increases  the  total  of  the 
charge  100  Ib.,  and  we  may  let  it  go  at  that,  submitting  to  a  little 
less  than  10%  lead  on  the  charge ;  or  we  may  decrease  the  silicious 
ore,  which  permits  us  to  lessen  the  quantity  of  flux.  Write  the  new 
figures  170  of  silicious  ore,  and  270  of  iron,  and  260  of  limestone, 
making  in  all  500  Ib.  for  the  three  constituents,  and,  after  neatly 
erasing  the  old  figures,  try  again.  The  amended  calculation  will 
be  nearly  correct,  or  if  not  so,  make  a  new  correction  with  the 
figures  erased  and  new  ones  inserted.  The  final  sheet  should  be  a 
fair  copy.  This  second  example  is  a  simple  one.  Frequently  the 
metallurgist  adds  other  items  to  the  charge,  and  methods,  other 
than  the  trial  one  here  given,  are  too  complicated. 

Since  the  atomic  weights  of  Fe  and  Mn  are  nearly  identical  (56 
to  55),  we  add  the  percentage  of  each  metal  to  obtain  the  equivalent 
Fe.  For  MgO  multiply  by  1.4,  and  for  BaO  by  0.4  obtain  the  equiva- 
lent CaO. 

131.     SAMPLING  AND  HANDLING  BASE-BULLION. 

The  practice  at  large  silver-lead  smelting-works  is  now  to  re-melt 
all  base-bullion  from  the  blast-furnace.  Sometimes  the  lead  is  taken 
in  molten  condition  to  the  re-melting  kettle  from  the  blast-furnace. 
.When  the  lead  is  melted  in  the  re-melting  kettle  it  is  carefully 
skimmed,  and  as  in  lead  refining,  the  cleaned  lead  is  molded  into 
bars.  The  skimming  or  dross,  containing  copper  and  other  impurity, 
is  returned  to  the  blast-furnace.  The  copper  there  enters  the  matte 
and  the  lead  again  goes  to  the  base-bullion.  While  the  lead  is 
being  molded  samples  are  taken  from  the  kettle  at  intervals,  and 
from  the  samples  the  assay-results  are  obtained.  The  bars  for  a 
40-ton  car-load,  800  in  number,  are  stamped  with  the  number  of 


OF    THE    COMMON    METALS.  391 

the  lot,  and  are  carefully  weighed,  20  at  a  time.     Careful  assays 
are  made  of  each  lot,  both  by  the  shipper  and  by  the  refiner. 

At  smaller  plants  the  punch  sample  is  taken  as  described  in 
tKe  chapter  on  sampling.  The  results  are  exact. 

132.     FLUE-DUST. 

A  blast-furnace  42  by  120  in.  takes  5000  cu.  ft.  of  air  per  min. 
when  in  full  operation.  The  escaping  gas  has  an  average  temperature 
of  150°C.,  and  we  calculate  that  the  velocity  of  the  gas  rising 
through  the  charge  is  at  least  5  ft.  per  sec.  Additional  air  enters 
the  feed-doors,  especially  at  the  time  of  charging,  and  particles  of 
20-mesh  size  may  be  carried  into  the  down-take  and  to  the  flue  that 
leads  to  the  tall  stack  producing  the  draft.  The  flue,  which  is 
common  to  all  the  blast-furnaces,  in  some  cases  is  made  hundreds 
of  feet  long  and  of  large  cross-section,  for  the  purpose  of  settling 
and  collecting  the  particles  called  'flue-dust.'  Flue-dust,  after 
suitable  preparation,  as  for  example  by  making  it  into  briquettes, 
can  be  re-melted.  It  amounts  to  0.8  to  15%  the  weight  of  the 
charge,  but  in  good  practice  a  near  approach  to  the  former  figure 
is  possible. 

A  little  of  the  lead  and  silver,  and  much  of  the  zinc  and  arsenic 
of  the  charge  are  volatilized.  The  volatilized  substance  in  part 
adheres  to  the  cool  surface  of  the  flue.  Eventually  it  flakes  off  and 
falls  to  the  bottom  of  the  flue  and  is  there  recovered.  Flue-dust, 
therefore,  is  composed  of:  (1)  dust,  carried  along  by  the  draft, 
and  (2)  lead  fume,  condensed  on  the  cool  surface  of  the  flue. 

All  this  material  is  not  recovered ;  the  finest  part  may  escape. 
An  analysis  of  flue-dust  at  Pueblo,  Colorado,  shows  PbO,  37.6%; 
ZnO,  5.3%  ;  Fe,03,  25%  ;  A12OS,  1.3%  ;  CaO  (from  the  limestone) 
5.3%;  SiO2,  8.6%;  S,  2.5%;  S03,  1.6%;  H20,  CO2,  and  C  (from 
the  coke)  11.2%.  When  arsenic  is  present  in  the  charge  a  part 
condenses  in  the  flue.  Where  in  outlying  regions,  the  lead  has  a 
low  value,  there  is  a  point  at  which,  in  order  to  eliminate  the 
arsenic,  it  is  better  to  permit  the  loss  of  some  of  the  lead. 

There  is  a  difference  in  the  quantity  of  flue-dust  made  by  a 
furnace  according  to  the  way  it  runs.  A  furnace  having  accretions 
or  scaffolds  in  the  shaft,  or  one  driven  with  a  large  volume  of  blast, 
or  one  fed  carelessly  so  that  the  gas  fails  to  ascend  evenly  through 
the  charge,  increases  the  production  of  flue-dust. 

The  carrying  power  varies  as  the  square  of  the  velocity,  or 
directly  as  the  draft-pressure,  therefore,  to  arrest  as  much  flue-dust 


392  THE    METALLURGY 

as  possible  in  the  main  dust-flue,  it  is  made  large  in  sectional  area 
thus  reducing  the  velocity  of  the  gas. 

Flues  have  been  made  of  sheet-metal,  but  metal  corrodes  because 
of  the  sulphuric  acid  that  is  found  in  the  gas  or  the  sulphates  that 
are  in  the  flue-dust.  Reinforced  concrete  has  been  used,  but  it 
also  is  somewhat  attacked.  Brick  remains  the  favorite  material 
for  such  construction.  A  flue  built  of  reinforced  concrete  on  the 
Monier  system  consists  of  an  arched  frame-work  of  angle-iron 
tied  by  longitudinal  bars  and  covered  by  wire  netting,  or  expanded- 
metal,  the  whole  being  plastered  inside  and  out  with  a  coating  of 
cement-concrete.  The  bottom  of  brick  flues  is  frequently  made  of 
sheet-steel  hoppers.  Since  these  are  continually  covered  by  the 
flue-dust,  they  are  protected  from  corrosion  and  last  a  long  time. 
A  flue  is  rectangular  in  cross  section  and  has  brick  walls,  a  low 
arched  top,  and  a  hopper-bottom,  as  above  described,  at  such  a 
height  as  to  leave  room  beneath  for  a  car  running  on  a  track  at 
the  ground  level.  The  car  can  be  set  under  any  hopper  and  the 
content  drawn  into  the  car  through  a  canvas  sleeve  fitted  over  the 
outlet  spout  to  avoid  dust. 

Other  methods,  such  as  spraying  water  upon  the  dust-laden  gas ; 
the  use  of  plates,  called  Freudenberg  plates,  hung  in  the  flue  parallel 
to  the  length;  Rosing  wires,  a  multitude  of  wires  hung  from  rods 
parallel  with  the  flue,  all  have  been  tried  with  a  degree  of  success, 
but  abandoned  in  favor  of  plain  flues  free  from  baffle-walls  or  other 
obstructions. 

The  two  important  considerations  in  the  construction  of  flues 
are,  first,  a  slow  and  general  movement  of  the  gas,  in  order  to 
settle  out  the  dust;  and  second,  a  large  cooling  surface  for  the 
condensation  of  the  lead-fume. 

133.     BAG-HOUSE. 

Fig.  163  is  a  transverse  section  of  the  bag-house  which  is  coming 
much  into  use  for  the  recovery  of  the  flue-dust  and  lead  fume  by 
filtering  the  gas  through  bags  that  leave  the  resultant  gas  colorless. 

It  is  a  building  40  ft.  wide,  200  ft.  or  more  long  and  45  ft.  high, 
divided  into  two  stories.  The  bags,  of  which  there  are  several 
hundred,  20  in.  diam.  by  35  ft.  long,  are  suspended  from  the  beams 
of  the  roof  by  cords  that,  at  the  same  time,  tie  the  mouths  of  the 
upper  ends.  The  second  floor,  14  ft.  from  the  ground,  is  pierced  with 
openings  18  in.  diam.  with  thimbles  over  which  the  lower  ends  of 
the  bags  are  slipped  and  tied.  The  flue-gas  is  drawn  in  by  a  suction- 
fan  and  delivered  into  the  lower  story,  beneath  the  second  floor. 


OF    THE    COMMON    METALS. 


393 


The  gas,  distending  the  bags,  becomes  filtered  and  passes  into  the 
second  story,  escaping  at  the  ventilator  in  the  roof.  The  building 
is  divided  by  several  cross-walls,  so  that  it  is  possible  to  cut  out  a 
section  when  it  is  desired  to  enter  it  to  make  repairs,  or  shake  the 
bags  one  by  one  as  is  done  once  or  twice  daily.  The  dust  falls  into 
the  lower  chamber  and  accumulates  there.  When  one  or  two  feet 
deep  it  is  ignited.  It  burns  of  itself,  becoming  agglomerated  and 


hingeddoor 


Bag  suspended  from  screw-eyes  tytfl°/8 
(or -better  Copper) w/ns 


Fig.    163.      BAG-HOUSE,   GLOBE   SMELTING   WORKS. 

in  a  condition  favorable  for  feeding  back  to  the  blast-furnace. 
Analysis  of  the  burned  dust  shows  it  to  contain  oxide  and  sulphate 
of  the  metals  as  follows:  Pb,  75%;  Zn,  3%  •  Fe,  0.5%;  As,  1.3%, 
and  Ag,  4  oz.  per  ton.  The  bag-house  has  been  introduced  because 
of  the  complaint  that  fumes  injure  vegetation,  and  also  because  of 
the  actual  saving  it  effects. 

134.     BRIQUETTING  FLUE-DUST. 

Flue-dust  can  be  wet  down  and  fed  back  to  the  blast-furnace. 
If  fed  a  little  at  a  time,  it  is  simply  carried  again  into  the  flue,  but 
while  wet,  in  occasional  large  charges,  it  may  be  fed  so  that  most 


394 


THE    METALLURGY 


of  it  is  carried  down  and  smelted.  The  effective  way  is  to  make 
it  into  briquettes  with  milk  of  lime  as  a  binder.  Fig.  164  represents 
a  plant  containing  a  White  briquetting-press  for  making  briquettes 
composed  of  flue-dust  and  milk  of  lime  to  which  is  added  fine 
roasted  ore.  At  the  right  in  the  figure,  shown  to  be  on  a  high  plat- 
form, is  a  pile  of  quicklime.  This  is  fed,  together  with  water,  into 
the  lime-mixer,  a  trough  divided  transversely  by  a  partition.  One 
compartment  is  shown  as  containing  the  lime  being  mixed  to  a 
thin  paste,  while  the  other  is  now  empty.  The  paste  is  drawn  from 
either  compartment  to  a  horizontal  double-shaft  pug-mill*  Each 


Fig.    164.      WHITE   BRIQUETTiNG   PRESS. 

shaft  is  provided  with  mixing  blades.  Flue-dust  from  the  pile  at 
the  front  of  the  lower  platform  is  shoveled  into  the  pug-mill  and 
thoroughly  mixed  with  the  milk  of  lime  by  the  revolving  blades  of 
the  pug-mill  which,  being  set  at  an  angle,  propel  it  to  the  discharge- 
opening  immediately  over  a  troughed  conveying-belt.  It  drops  into 
the  hopper  of  a  six-mold  briquetting-press  where  it  is  made  into 
briquettes  that  drop  upon  a  flat  conveying-belt,  delivering  them 
to  a  pile.  The  briquettes  may  be  used  at  the  blast-furnace  freshly 
made,  but  the  usual  plan  is  to  dry  them,  as  clay-bricks  are  dried. 

At  times  the  briquetting  is  omitted,  and  the  pug-mill  mixture 
is  wheeled  to  a  drying-floor,  or  is  distributed  evenly  upon  one  of 
the  ore-beds.  By  the  time  the  bed  is  used  the  mixture  has  set  and 
become  a  hard  mass  capable  of  withstanding  handling  without  being 
broken. 

The  lead-smelting  charge  generally  contains  copper,  and  the 
copper  accumulates  in  the  matte.  Since  matte  is  roasted  and 


OF    THE    COMMON    METALS.  395 

135.     LEAD-COPPER  MATTE. 

returned  to  the  blast-furnace,  the  content  in  copper  gradually 
increases.  When  increased  to  12%,  the  copper  matte  is  again 
roasted  and  treated  in  a  separate  blast-furnace,  with  silicious  ore 
and  oxidized  copper  ore,  to  produce  a  matte  of  40%  copper,  called 
'shipping  matte'  because  it  often  is  shipped  to  a  copper  works  to 
be  treated  for  the  copper. 

A  satisfactory  way  of  treating  matte,  where  it  is  wished  to 
produce  a  more  finished  product,  is  to  crush  it  to  a  4-mesh  size  and 
roast  in  a  hand-roaster.  It  is  next  sent  to  a  blast-furnace  and 
again  smelted  with  silicious  and  oxidized  ores  of  high  grade  in 
copper.  There  results  a  matte  65%  Cu  and  a  proportion  of  bottoms, 
the  result  of  the  reduction  of  the  copper  from  the  matte.  The 
bottoms  are  charged  into  a  reverberatory  furnace  through  the  side- 
doors,  and  the  coarse-broken  matte  is  put  on  top.  The  doors  are 
closed,  and  the  charge  is  fired  with  an  oxidizing  flame,  as  in  the 
Welsh  process  of  'roasting.'  The  charge  having  melted,  a  reaction 
of  the  cuprous  oxide  on  the  cuprous  sulphide  takes  place,  as 
described  in  the  Welsh  process  of  making  blister  copper;  and  the 
charge  becomes  reduced  to  an  impure  copper  containing  arsenic, 
bismuth,  and  antimony,  as  well  as  the  gold  and  silver  that  were 
contained  in  the  matte.  The  copper  is  then  poled  to  reduce  the 
cuprous  oxide,  and  ladled  into  anode-molds.  The  anodes  are  sent 
to  an  electrolytic  copper  refinery  for  final  treatment. 

136.  COST  OF  TREATMENT  OF  LEAD  ORES. 

To  illustrate  the  method  of  calculating  the  actual  cost  of  treating 
an  ore,  as  in  the  Colorado  or  Utah  silver-lead  smelting  practice, 
we  take  the  case  of  a  so-called  neutral  ore  (SiO2  equal  to  Fe).  The 
ore  is  assumed  to  be  oxidized,  to  contain  less  than  5%  sulphur,  and 
at  least  10%  lead.  It  is  to  be  treated  at  a  works  having  an  output 
of  400  tons  of  charge  daily. 

The  cost  of  treating  a  ton  of  charge,  and  of  treating  a  ton  of 
the  ore  including  the  flux,  is  as  follows: 

Cost  per  ton 

of  charge.  of  ore. 

Labor    $1.10  $1.54 

General  expense,  assaying,  and  management....  0.20  0.27 

Fuel    for    power 0.07  0.10 

Coke   (15%  of  the  charge)   at  $6.50  per  ton 0.97  1.36 

Interest,   improvement-fund,   and   repairs 0.26  0.36 

Limestone    (0.3   ton   at   $1.25) 0.37 

Iron  ore  (0.1  ton  at  $5 ) '. ' 0.50 

$2.60  $4.50 


396  THE    METALLURGY 

The  figures  in  the  second  column  are  obtained  by  multiplying 
the  total  weight  of  the  charge,  1.4  tons,  by  the  cost  of  each  item  per 
ton  of  material,  and  then  adding  the  cost  of  the  flux.  This 
corresponds  to  the  figure  above  obtained,  and  may  be  stated  again 
as  follows : 

1.4  tons  of  material  (1  ton  ore,  0.3  ton  lime- 
stone, and  0.1  ton  of  iron  ore)  at  $2.60  per 

ton  of  charge  for  smelting $3.63 

Cost  of  fluxes,  0.3  ton  of  limestone  at  $1.25.  .  .   0.37 
Cost  of  fluxes,  0.1  ton  of  iron  ore  at  $5.00 0.50 


$4.50 

In  case  the  ore  contains  sulphur  in  quantity  to  require  roasting, 
$2  should  be  added  for  the  treatment.  For  sulphur  over  5%  and  up 
to  10%,  add  30c.  per  unit  to  cover  the  expense  of  iron  ore  for 
disposing  of  the  extra  sulphur,  and  roasting  the  matte  made  by  it. 

The  application  of  this  method  of  calculating  costs  may  be 
illustrated  with  a  silicious  ore.  We  make  a  charge-calculation  and 
find  that  for  one  ton  of  ore  we  need  1800  Ib.  iron  ore  and  1450  Ib. 
limestone.  Thus  we  have  : 

5,250  Ib.  or  2.625    tons    of   material    of   the 

charge  at  $2.60  per  ton $6.82 

1,800  Ib.  or  0.9  ton  ore  at  $5  per  ton 4.50 

1,450  Ib.  or  0.725  ton  of  limestone  at  $1.25     1.08 


$12.40 

This  $12.40  represents  the  actual  cost  of  smelting  a  ton  of  silicious 
ore  when  the  fluxes  are  paid  for  outright,  and  when  run  in  a  furnace 
also  with  suitable  lead-bearing  ores.  In  general,  the  lacking  iron 
for  the  charge  is  made  up  in  part  by  the  excess  of  iron  over  that 
needed  for  the  sulphur  and  silica  contained  in  the  roasted  iron- 
sulphide,  and  in  the  other  irony  ores. 

A  ton  of  the  above  silicious  ore  makes  1.5  tons  of  slag  which, 
containing  0.6%  Pb  and  0.6  oz.  Ag  per  ton  causes  a  loss  in  these 
metals  of  $1.26  per  ton.  The  self-fluxing  ore  above  cited  produces 
1250  Ib.  slag  and  carries  off  only  $0.52.  If  we  were  treating  this 
ore  alone,  of  100  tons  treated  daily  but  38  tons  would  be  ore,  and 
from  this  all  the  profits  of  smelting  would  have  to  be  obtained. 
In  the  case  of  the  self-fluxing  ore,  71  tons  are  metal  bearing.  Thus 
the  room  taken  by  the  fluxes,  called  'displacement,'  is  taken  into 
account  in  figuring  the  profit  of  operation. 


PART  VIM.     ZINC 


PART  VIII.     ZINC. 

137.     PROPERTIES  OF  ZINC. 

Zinc  is  a  white,  brittle  metal  of  7.1  to  7.2  specific  gravity. 
Commercial  zinc,  called  spelter,  contains  lead,  iron,  and  cadmium 
as  common  impurities.  Lead  is  found  in  spelter  to  the  extent  of 
2  to  3%,  but  the  amount  can  be  reduced  to  1%  by  refining.  A  small 
amount  increases  ductility,  but  it  is  injurious  in  the  better  grades 
of  brass.  Iron  may  be  present  up  to  0.05%.  Cadmium  seldom 
occurs  in  spelter  and  has  no  deleterious  effect,  except  when  the 
metal  is  used  for  making  zinc-white.  It  then  discolors  the  product. 

138.     ZINC  ORES. 

The  principal  ores  of  zinc  are  blende  and  calamine.  In  New 
Jersey  occur  deposits  of  franklinite.  The  minerals  seldom  occur 
pure ;  besides  the  earthy  gangue,  and  sulphides  of  iron,  lead,  and 
copper,  ffiQy  are  fl"Q/in^ntly  centaminntcd  with  impuritiryh  combined. 
chemically,  nnpnmnllj  -irun>  which  cannot  be  separated  by  ore- 
dressing  methods. 

Blende,  or  sphalerite,  when  of  a  yellow  color  as  in  the  ore  of 
the  Joplin  district  in  Missouri,  is  called  rosin-blende.  When  dark 
in  color,  due  to  chemically-contained  iron  as  in  the  ore  of  the  Rocky 
Mountain  States,  it  is  called  black-jack.  It  is  from  blende  that 
most  of  the  spelter  of  commerce  is  extracted.  It  needs  roasting 
before  it  can  be  retorted  or  smelted  for  extracting  the  zinc. 

Calamine  is  a  term  applied  commercially  to  the  carbonate 
(smithsonite)  and  to  the  hydrous-silicate  of  zinc.  It  is  an  oxidized 
or  sulphur-free  ore  that  needs  no  preliminary  roasting  before  smelt- 
ing. On  being  heated  in  the  retort,  the  CO2  of  the  carbonate  is 
expelled,  leaving  zinc  oxide.  Willemite,  the  anhydrous  silicate  occur- 
ing  with  franklinite  in  New  Jersey,  mixed  with  coal  is  decom- 
posed at  the  high  temperature  of  the  retort,  yielding  zinc.  Never- 
theless it  is  generally  found  advantageous  to  calcine  calamine  for 
the  purpose  of  driving  off  C02  and  water,  which  are  undesirable  in 
retorting  because  of  their  oxidizing  action  on  zinc-vapor.  However, 


400  THE    METALLURGY 

the  preliminary  calcination  is  often  omitted,  but  when  performed,  it 
is  done  in  kilns  much  like  those  in  which  lime  is  burned. 

139.     METALLURGY  OF  ZINC. 

In  outline,  the  metallurgy  of  zinc  consists  in  grinding  the  ore 
(generally  blende)  and  roasting  to  convert  it  into  zinc  oxide,  then 
charging  the  roasted  ore,  intimately  mixed  with  fine  coal,  into 
horizontal,  cylindrical,  clay  retorts,  heated  to  a  white  heat,  where 
the  zinc,  reduced  by  the  coal,  volatilizes,  and  the  vapor,  entering 
the  cool,  tapering,  clay  extension  of  the  retort  (called  the  condenser), 
collects  there.  As  it  accumulates  it  is  tapped  into  a  ladle  from  time 
to  time,  skimmed,  and  cast  in  molds.  When  distillation  is  complete 
the  condenser  is  removed  and  the  content  of  the  retort  taken  out 
and  generally  thrown  away.  The  cycle  of  operations  takes  24 
hours. 

140.     ROASTING  BLENDE. 

The  aim  is  to  dead-roast  the  ore,  generally  to  1%  sulphur  or 
less.  For  every  1%  sulphur  remaining  in  the  roasted  ore,  2%  zinc 
is  held  back  in  the  retort-residue  after  distillation. 

The  ore  is  ground  to  8-mesh  size  then  slowly  and  carefully 
roasted  with  frequent  stirring,  finishing  the  roast  at  a  high  tem- 
perature to  decompose  the  zinc  sulphate  formed  at  the  lower  tem- 
perature. The  ore  is  generally  in  the  form  of  concentrate,  still 
containing  a  little  gangue,  galena,  and  pyrite.  To  remove  the 
final  1%  of  sulphur  would  require  a  long  time  and  would  not  be 
commercially  profitable.  The  ore  is  accordingly  considered  to  be 
finished  when  it  contains  no  more  than  that  amount  of 'sulphur. 

141.     CHEMISTRY  OF  ROASTING  ZINC  ORES. 
We  have,  in  the  roasting  of  blende,  the  following  reactions : 

(1)  ZnS  +  40  =  ZnO  +  S03 

43,000  86,400  71,000  =  +  114,400  cal. 

(2)  2ZnS  +  70  =  ZnO  +  ZnS04  +  S02 

2x43,000  86,400  230,000     71,000  ===  +  301,400  cal. 

Thus  in  an  oxidizing  flame,  blende  is  roasted  to  oxide  and 
sulphate,  both  reactions  being  exothermic.  As  indicated  in  the 
reactions  given  in  the  chapter  on  roasting,  pyrite  or  chalcopyrite 
assists  in  the  reactions.  At  a  cherry-red  heat  the  zinc  sulphate  is 
decomposed  into  basic  sulphate  (3ZnO,ZnS04)  thus: 

(3)  4ZnS04  =  3ZnO,ZnS04  +  3SO3 

The  basic  sulphate,  exposed  to  a  bright-red  heat  for  a  time,  reacts 
thus: 


OF   THE    COMMON    METALS.  401 

(4)  3ZnO,ZnS04  =  4ZnO  +  SO3 

Finally  zinc  oxide  is  obtained  and  the  sulphuric  anhydride  is 
eliminated.  At  a  high  temperature  (900°C.)  the  latter  in  part 
decomposes  into  sulphurous  anhydride  and  oxygen  as  follows : 

(5)  S03  =  S02  +  O 

When  limestone  or  calcite  is  present  it  is  converted  in  part  to 
sulphate  and  in  part  to  oxide.  Galena  also  roasts  to  a  sulphate, 
and  tends  to  envelop  particles  of  blende,  and  to  prevent  their 
roasting.  Much  of  the  blende  from  Leadville,  Colorado,  contains 
silver,  and  it  consequently  often  pays  to  treat  the  retort-residues 
after  the  zinc  has  been  removed. 

There  is  a  loss  of  silver  in  roasting  that  may  be  given  at  10%, 
and  also  a  loss  of  zinc  as  dust  and  volatilization  at  the  final  high 
temperature  that  may  be  reckoned  at  2  per  cent. 

142.     ROASTING  FURNACES. 

The  roasting  of  blende  has  been  performed  in  hand-rabbled 
reverberatory  furnaces  as  well  as  in  a  great  variety  of  mechanical 
furnaces.  These  are  described  in  the  chapter  on  roasting.  The 
latter  are  gradually  supplanting  the  former  because  of  the  saving 
of  labor.  It  should  be  noted,  however,  that  the  wear  is  great  on 
mechanical  furnaces  that  have  iron-work  exposed  to  the  heat  because 
of  the  high  final  heat  needed  in  blende-roasting,  and  consequently 
the  types  of  furnace  have  been  preferred  where  the  rabble  is  exposed 
but  a  short  time  to  the  action  of  the  fire,  and  where  iron  parts  are 
not  exposed  or  can  be  water-cooled. 

Thus,  the  Brown  horseshoe  furnace,  where  the  rabble  is  drawn 
through  a  circular  hearth,  then  allowed  to  cool,  or  the  Wethey 
furnace,  where  the  rabble  is  exposed  to  the  fire  but  half  of  the 
time  and  the  moving  iron  parts  are  outside  the  furnace,  have  been 
successfully  used  in  blende-roasting.  Of  the  recent  types,  the 
Hegeler  furnace  as  used  by  the  Matthiessen  &  Hegeler  Co.,  La  Salle, 
Illinois,  and  by  the  U.  S.  Zinc  Smelting  Co.,  Pueblo,  Colorado,  has 
proved  most  successful  for  the  above  reasons.  It  is  a  multiple- 
hearth  furnace  closed  by  swinging  sheet-iron  doors  at  the  ends,  and 
stirred  by  rabbles  drawn  quickly  through  the  furnace  by  means  of 
rake  rods,  so  that  the  parts  are  outside  the  furnace  most  of  the  time, 
and  no  iron  parts,  except  the  end  swinging  doors,  are  affected  by 
the  fire.  The  hearths  being  superimposed  make  a  compact  furnace, 
and  the  radiation  is  greatly  lessened,  so  that  there  is  economy  of 
fuel.  A  75-ft.  hearth  Hegeler  furnace  roasts  48  tons  of  blende  in 


402  THE    METALLURGY 

24  hours,  yielding  40  tons  of  roasted  ore.  The  raw  ore  contains 
30%  sulphur,  and  the  final  roast  1.2%.  The  consumption  of  coal  is 
approximately  20%  of  raw  ore. 

143.     THE  SMELTING  OR  DISTILLATION  OF  ROASTED  ZINC 

ORES. 

The  recovery  of  zinc  from  the  ore  consists  in  distillation  of  the 
roasted  ore  in  refractory  clay  retorts  after  intimately  mixing  it 
with  40  to  50%  the  weight  of  fine  coal.  The  whole  is  brought  to  a 
white  heat  which  is  maintained  during  an  entire  day. 

Before  ore  and  coal  are  charged  into  the  hot  retort  the  mixture 
is  moistened  for  convenience  in  charging,  the  water  promptly  being 
driven  off  by  the  heat.  The  light  hydrocarbons  of  the  coal  come 
away  next;  then  the  iron  oxide  is  reduced  to  protoxide  and  part 
of  it  to  a  porous  iron  or  iron  sponge.  The  final  reaction  is  the 
reduction  of  the  zinc  oxide  of  the  ore  by  the  carbon  to  metallic 
zinc.  The  reaction  commences  at  1060°C.,  but  practically  a 
temperature  of  1300°C.  is  reached. 

In  the  United  States  the  Belgian  system  of  retorting  is  chiefly 
used.  The  retorts  4  ft.  long  by  81/-?  in-  inside  diameter  are  set 
horizontally  in  the  hot  furnace.  The  retort  and  charge  being  poor 
conductors  of  heat,  1.5  to  2  hours  is  necessary  for  the  heat  to  pass 
from  the  hot  exterior  to  the  cold  core.  For  ore  high  in  iron,  the 
furnace  in  that  time  becomes  heated  to  a  uniform  temperature 
(1020°C.)  and  the  iron  oxide  is  reduced.  If  the  heat  rises  above 
this  point  rapidly,  the  reduction  of  the  oxides  of  iron  and  zinc 
occurs  simultaneously  and  the  CO  and  C02  gas  resulting  from  the 
reduction  of  the  iron  oxide  tends  to  sweep  the  zinc-vapor  through  the 
condenser,  and  cause  it  to  burn  at  the  mouth.  There  is  a  difference 
in  the  reductibility  of  iron-bearing  blende,  so  that  a  hard  ore, 
difficult  to  roast,  is  naturally  more  difficult  to  reduce  than  a  porous 
one  that  has  been  roasted  at  a  low  temperature.  If,  as  sometimes 
happens,  a  zinc  ore  is  roasted  at  so  high  a  temperature  as  to  form 
an  incipient  slag,  especially  an  iron  silicate,  the  reduction  of  the 
ore  is  retarded.  The  iron  silicate  collecting  at  the  bottom  of  the 
retort  attacks  other  bases  and  quickly  cuts  a  hole  through  the 
retort.  The  iron  sponge,  where  formed,  is  not,  however,  detrimental, 
but  where  sulphur  is  still  left  in  the  ore,  it  may  combine  with  it  to 
form  iron  sulphide  thus: 

(6)     ZnS  +  Fe  =  FeS  +  Zn  " 

The  zinc  is  released,  it  is  true,  but  the  iron  sulphide  formed  at 
1100°  to  1200°C.  is  corrosive  in  action  upon  the  retort. 


OF    THE    COMMON    METALS.  403 

When  the  content  of  the  retort  has  been  brought  to  the  reduction 
•temperature  for  zinc  oxide  we  have : 

(7)  ZnO  +  C  —  Zn  +  CO 

86,000  29,000=  -57,000  cal. 

Thus  the  zinc  oxide  is  reduced  by  carbon  to  zinc  vapor  with  the 
formation  of  carbon  monoxide.  The  carbon  monoxide  escaping-  from 
the  mouth  of  the  condenser,  burns  with  a  characteristic  light  blue 
flame,  masked,  however,  by  a  small  amount  of  zinc-vapor  that 
escapes  and  burns  at  the  same  time. 

Any  undecomposed  ZnS,  left  in  the  roasted  ore,  remains  as  such 
unless  iron  sponge  is  formed  to  reduce  it  as  above  described.  Zinc 
sulphate  is  reduced  to  sulphide  according  to  the  following  reaction : 

(8)  ZnSO4  -f  2C  =  ZnS  +  2C02 

In  the  presence  of  an  excess  of  carbon  the  carbon  dioxide  may  be 
changed  to  carbon  monoxide  thus : 

(9)  CO2  +  C  =  2CO 

If,  however,  any  dioxide  remains  unreduced,  it  tends  to  oxidize  the 
zinc  vapor  back  to  zinc  oxide,  which  then  escapes  from  the  mouth 
of  the  condenser  and  is  lost. 

Lead  sulphate,  or  oxide  in  roasted  ore  is  reduced  to  metallic 
form  and  remains  in  part  in  the  residue  of  the  retorts  and  in  part 
is  volatilized  and  condensed  with  the  zinc,  contaminating  the  spelter. 
Cadmium  oxide  is  present  to  the  extent  of  a  few  tenths  of  a  per 
cent  in  certain  zinc  ores.  It  is  more  volatile  and  easily  reduced 
than  zinc  oxide,  and  condenses  with  the  first  of  the  zinc.  It  has 
been  known  to  be  present  in  spelter  to  the  extent  of  0.5%,  but 
usually  there  is  but  a  trace.  It  has  not  been  found  injurious  to 
spelter.  Silver  and  gold  remain  in  the  residue  after  retorting.  In 
Missouri,  where  there  is  little  or  no  precious  metal,  the  residue  is 
thrown  away,  but  Colorado  ores,  carrying  silver  and  gold,  yield 
residues  suited  to  further  treatment,  in  the  blast-furnace,  for  the 
extraction  of  the  metals. 

To  give  an  idea  of  the  result  to  be  expected  from  roasting  and 
retorting  an  iron-bearing  blende  we  take  the  following:  A  charge 
consisted  of  roasted  ore  containing  Zn  44.8%,  Pb  4.7%,  Fe  18.0%, 
CaO  1.1%,  MgO  0.7%,  SiO,  6.0%,  and  S  2.9%.  It  was  retorted  in  a 
charge  of  60%)  ore  and  40%  fuel  for  24  hours,  and  there  resulted 
35%  of  residue  computed  on  the  total  charge.  The  residue  contained 
0.73%  Zn.  24.2%  Fe,  and  4.07%  S,  and  of  the  total  zinc,^95%  was 
/rotaincfl.  Where  the  same  ore  was  imperfectly  roasted  and 
contained  7.7%  S,  it  was  found  that  6.7%  Zn  was  lost  in  the  residue. 


404 


THE    METALLURGY 


This  shows  that  within  commercially  profitable  limits,  too  much 
care  cannot  be  exercised  to  effect  a  good  roast.  Now  according  to 
the  rule  stated  in  the  chemistry  of  roasting,  the  loss  should  have 
been  15.4%  Zn.  We  must  accordingly  infer  that  the  greater  extrac- 
tion of  zinc  is  due  to  the  reducing  effect  of  the  iron  present  reacting 
on  sulphates  and  sulphides  containing  zinc. 

144.     THE   ZINC  FURNACE. 

Fig.  165  and  166  represent,  in  transverse  section  and  in  front 
elevation,  a  Western  zinc-smelting  furnace  after  the  Belgian  system, 


Fig.   165.      ZINC-SMELTING 
FURNACE    (SECTION). 


Fig.   166. 


ZINC-SMELTING  FURNACE 
(ELEVATION). 


for  the  distillation  of  roasted  blende.  It  contains  265  retorts,  or 
128  on  each  side.  As  shown  in  the  transverse  section,  there  are 
8  rows  of  retorts,  the  rear  ends  of  which  rest  on  the  middle  wall  of 
the  furnace.  The  front  ends  rest  on  the  front  walls.  Beneath 
the  retorts  are  the  fire-places,  each  18  in.  wide  by  8  ft.  long,  two 
under  each  bank  of  retorts.  These  are  fired  from  the  four  end  doors. 
Bituminous  coal  is  used,  the  fire-bed  being  4  ft.  deep,  so  that  the 
coal  is  burned  with  a  long  flame  that  completely  fills  the  interior 
of  the  furnace  and  surrounds  the  retorts.  The  products  of  com- 
bustion pass  through  outlet-ports  in  the  roof  to  one  of  the  two  stacks, 
one  at  each  end  of  the  block  or  'massive'  of  the  furnace.  Beneath 


OF    THE    COMMON    METALS.  405 

the  grates  ample  room  is  left  for  a  man  to  work,  and  with  a  long 
bar,  to  'grate'  or  clean  the  fires,  removing  the  clinker  and  ashes 
as  they  form.  Two  or  three  fire-bars,  each  2  in.  square,  sustain 
the  fire.  This  underground  passage  also  serves  for  a  tram-car  in 
which  the  ashes  are  carried  away,  and  the  residue  from  the  retorts 
removed.  The  end  fire-doors  have  the  sills  at  the  ground  level, 
2  ft.  below  the  lowest  retorts.  In  Fig.  141  and  142  only  the  three 
lower  rows  of  retorts  shown  are  provided  with  condensers.  In  full 
action  all  are  so  provided.  The  block  proper  is  14  ft.  high,  24  ft. 
long,  and  12  ft.  from  face  to  face.  The  stacks  or  chimneys  are 
45  ft.  high,  one  for  each  side  of  the  block. 


Fig.    107.      ZINC-SMELTING   RETORTS. 

Fig.  167  shows  in  detail  a  zinc-retort  in  place  in  the  furnace. 
It  is  made  of  fire-clay,  4  ft.  long  by  8.5  in.  diam.  and  with  walls 
1.25  in.  thick.  In  the  figure  a  is  the  retort,  which  rests  on  a  ledge  on 
the  rear  wall  of  the  furnace  and  extending  just  through  the  thin 
(41/2-in.)  front  wall.  The  wall  is  held  by  buck-staves  c  which 
carry  the  tiles  upon  which  the  retorts  rest.  The  whole  is  firmly 
bound  together  with  tie-rods.  When  the  retort  has  been  charged, 
the  clay  condenser  &  is  set  in  place,  in  which  the  zinc  vapor  issuing 
from  the  retort  is  to  condense.  As  seen  in  the  front  view,  the 
space  between  two  buck-staves  is  divided  by  shelves  which  form 
'pigeon-holes,'  each  of  which  contains  two  retorts.  The  retorts 
having  been  set  in  place,  the  opening  around  them  is  bricked  up 
with  pieces  of  brick  and  with  clay.  When  a  retort  becomes  cracked 
or  otherwise  useless,  it  can  be  readily  removed  by  breaking  away 
the  temporary  wall  and  another  retort  set  in  the  place  without 
disturbing  the  adjacent  retorts. 

145.     OPERATION  OF  THE  FURNACE. 

The  roasted  ore  is  thoroughly  mixed  with  fine  coal  upon  the 
floor  of  the  retort-house  with  shovels,  or  better  in  a  horizontal 
pug-mill,  using  water  to  dampen  it.  The  charge  for  a  retort  con- 
sists of  60  Ib.  ore  and  40  Ib.  coal,  and  is  skilfully  and  rapidly 


406  THE    METALLURGY 

thrown  into  the  retort,  empty  from  the  last  charge  and  already 
red  hot,  using  the  special  scoop  shovel.  Fig.  168,  and  filling  it 
clear  to  the  back.  An  iron  rod  is  now  run  in  along  the  top  of 
the  charge  to  form  a  channel  for  the  escape  of  the  gas,  and  the 
condenser  _&,  Fig.  167,  a  conical  clay  tube  2  ft.  long  is  adjusted 
with,  a  piece  of  brick  to  sustain  it  in  the  exact  position.  The  j  j'mt 


*• 


a 


6 


Fig.    168.      CHARGE   SCOOP. 

between  the  condenser  and  the  retort  is  luted  with  a  loamy  clay. 
To  make  the  joint  tight,  a  crescent-shaped  'stamper,'  Fig.  169,  is 
used,  by  which  the  loam  is  compressed  and  the  joint  made  tight 
around  the  condenser.  The  tool,  before  using,  is  heated  to  dull 
redness  in  one  of  the  lower  rows  of  retorts,  called  'cannons.' 
These  cannons  are  not  used  in  retorting,  being  left  in  place  to 
modify  the  direct  intense  heat  of  the  fire  below. 


Fig.    169.      STAMPER. 

The  retorts  are  now  charged  and  brought  up  to  a  white  heat 
(1300°C.),  the  temperature  in  the  retort  being  about  100°  lower 
than  that  outside.  As  the  charge  becomes  heated,  the  moisture 
and  volatile  gas  come  away,  reduction  of  the  iron  takes  place,  and 
finally,  the  zinc  oxide  is  reduced  to  zinc  vapor,  coming  away 
together  with  the  CO  formed  at  the  same  time,  and  the  reactions 
tt/uM/v^take  place  as  already  shown.  The  zinc  vapor,  escaping  from  the 
charge  by  its  own  tension,  enters  the  cool  J?sfcsit  and  liquefies 
by  being  cooled  below  the  boiling  point.  In  practice  the  tem- 
perature in  the  condenser  is  550°C.  The  end,  or  mouth,  of  the 
condenser,  which  is  about  21/4  in.  diam.  is  loosely  plastered  with  a 
handful  of  the  charge-mixture.  From  time  to  time  (four  times  in 


OF    THE    COMMON    METALS.  407 

the  24  hours)  the  accumulated  zinc  is  scraped  into  a  ladle  held 
below  the  mouth,  with  a  button-headed  scraper.  The  opening  is 
at  once  loosely  plastered  up  again.  Toward  the  end  of  the  distil- 
lation the  firing  is  more  vigorous  to  remove  the  last  trace  of  metal. 
The  charge  is  put  into  the  retorts  early  each  morning  and  is  under 
the  action  of  the  fire  nearly  24  hours.  After  this  time,  the  final 
zinc  is  removed,  and  the  men  take  down  the  condensers,  breaking 
the  joint  formed  between  the  retort  and  condenser,  and  remove 
each  condenser  to  be  carefully  scraped  out  and  again  used. 

The  front  layer  of  residue  in  the  retort,  which  retains  zinc, 
is  again  retorted.  The  rest  of  the  content  of  the  retort  is  scraped 
out  with  a  suitable  scraper  or  rabble.  In  Missouri  it  is  often 
done  by  inserting  a  steam  pipe  to  the  back  of  the  retort  and  blow- 
ing out  the  content.  The  residue  is  rejected  if  it  contain  no  precious 
metal,  otherwise  it  is  smelted.  The  operation  of  charging,  as 
already  described,  is  at  once  begun,  except  when  retorts  are 
broken  that  have  to  be  replaced. 

Replacing  retorts. — The  average  service  of  a  retort,  in  good 
practice,  is  40  charges.  In  the  case  of  the  block  above  described 
it  would  be  necessary  to  replace  six  daily,  though  this  will  vary 
greatly  from  the  average  given.  Some  retorts  break  after  a  charge 
or  two ;  others  have  a  long  life.  Thus  a  retort  may  be  corroded  by 
iron  that  eventually  perforates  it.  A  retort  may  have  a  crack,  or 
may  become  cracked,  so  that  the  zinc  vapor  escapes ;  and  it  must 
be  replaced  by  a  fresh  one.  Such  an  accident  is  indicated  by  the 
disappearance  of  the  slight  flame  and  smoke  that  generally  issue 
from  the  condenser,  there  then  being  an  inward  current,  due  to 
the  chimney-draft  drawing  the  air  through  a  crack  of  the  broken 
retort.  In  such  a  case  the  charge  may  as  well  at  once  be  removed 
from  the  retort. 

Retorts  are  not  set  in  place  when  cold,  but  are  heated  red-hot 
in  a  coal-fired  annealing-furnace.  A  supply  of  them  is  ready  in 
this  furnace  each  morning  for  use.  The  old  retort  is  first  removed 
by  breaking  out  the  adjoining  wall  of  the  'pigeon-hole.'  The  new 
one  when  needed  is  brought  to  the  zinc-furnace  by  three  men,  and 
at  once  put  in  place.  The  work  is  hot  and  arduous.  The  fresh 
retort,  having  been  set,  is  quickly  bricked  in,  and  is  ready  for 
charging. 

A  charge  of  60  Ib.  of  60%  zinc  ore  if  well  roasted  contains 
45  Ib.  zinc  oxide  or  36  Ib.  zinc,  and  yields  but  85  to  90%  of  the 
zinc.  The  residue  of  a  100-lb.  charge  containing  ZnO  that 


408  THE    METALLURGY 

persistently   remains   in   the   residue,   and   ZnS   that  is   completely 
reduced  contains: 

Pounds. 

Zinc  sulphide  and  zinc  oxide 2.5 

Gangue  and  ash  (from  ore  and  burned  coal) .  .25.0 
Unburned  coal    (two-thirds  having  been  used 
in  reduction)    13.0 


Total    . ...40.5 

The  reduction  in  the  total  weight   (100  Ib.)   is  thus  59.5  per  cent. 

146.  MANUFACTURE  OF  RETORTS  AND  CONDENSERS. 

Retorts  to  withstand  the  high  temperature  and  corrosive  action 
of  the  charge  are  made  of  the  most  compact  and  durable  material. 
The  material  consists  of  a  mixture  of  'chamotte'  'grog'  or  'cement/ 
composed  of  burned  fire-clay,  or  made  from  old  broken  retorts,  fire- 
brick, or  tile  free  from  slag.  It  is  ground  to  3  or  4-mesh  size  and 
mixed  with  an  equal  amount  of  raw  fire-clay.  The  mixing  is  done 
in  a  pug-mill  using  10%  water  to  form  a  stiff  mud  or  'adobe,' 
which  is  allowed  to  stand  2  to  4  weeks  covered  with  wet  sacking  to 
season  and  to  develop  the  plasticity.  It  is  again  put  through  the 
pug-mill,  and  finally  made  into  retorts  in  a  hydraulic  retort-making 
machine  under  the  pressure  of  3000  Ib.  per  sq.  in.  A  machine  of  this 
kind  makes  8  retorts  per  hour  at  a  calculated  cost  of  50c.  per 
retort. 

The  success  of  the  retorting  operation  depends  upon  the 
durability  of  the  retorts,  and  thus  upon  the  composition  and 
manufacture.  Since  the  bases  in  the  charge  attack  the  retort,  we 
either  must  exclude  them  by  selection  of  ores,  or  use  less  silicious 
material  in  the  retort  mixture.  A  retort  carrying  an  excess  of 
silica  is  more  corroded  than  one  containing  a  large  proportion  of 
clay  or  aluminous  material.  On  the  other  hand,  if  we  increase  the 
proportion  of  clay,  the  retort  is  liable  to  shrink  and  crack  in 
making,  and  it  is  not  so  infusible  nor  as  stiff  as  one  made  of  a 
highly  silicious  mixture.  Good  air-dried  clay  for  retorts  consists 
of  30%  A1203,  50%  Si02,  15%  combined  water,  and  5%  bases. 

The  requirements  for  a  good  retort  are :  that  it  be  made  of 
refractory  material  to  resist  the  intense  heat,  that  it  be  not  quickly 
corroded  by  impurities,  that  it  be  dense  so  as  to  prevent  the  zinc- 
vapor  from  penetrating  it,  that  it  be  strong  to  keep  the  shape 
when  loaded  with  the  charge  of  ore  and  coal  and  heated  to  a 


OF    THE    COMMON    METALS.  409 

white  heat  (1300  to  1400°C.),  and  that  it  have  thin  walls,  not  to 
exceed  l1/^  in.,  to  permit  the  heat  to  be  readily  transmitted  to  the 
contents.  It  is  evident  that  upon  increasing  the  diameter  of  the 
retort  the  heat  enters  the  charge  more  slowly,  and  hence  we  limit 
the  diameter.  The  distance  of  4  ft.  between  the  supports  is  not 
exceeded  because  the  retort  would  sag  under  the  load  at  the  high 
temperature. 

Condensers. — Condensers  are  made  of  less  refractory  clay  than 
retorts.  They  are  not  subjected  to  high  temperature  but  must 
withstand  much  handling  and  severe  treatment.  They  last  8  to  12 
days,  and  cost  3  to  4c.  each. 

Drying  the  retorts. — The  finished  retorts  are  removed  to  a  drying 
house  and  dried  6  to  25  weeks.  Long  drying  gives  better  quality. 
They  are  first  put  into  a  room  of  ordinary  temperature  and  kept  10 
to  15  days  until  they  can  be  safely  handled.  They  then  are  taken 
to  a  hot-room  and  further  dried  at  a  temperature  30  to  35  °C.  until 
all  the  mechanically  contained  water  is  eliminated. 

Annealing. — Before  a  retort  can  be  put  into  the  furnace  it  must 
be  carefully  annealed  to  remove  the  chemically  combined  water. 
Accordingly  retorts  are  put  into  the  annealing  furnace  and 
gradually  heated,  so  as  to  be  ready  and  at  the  required  temperature 
on  the  morning  of  the  day  that  they  are  to  be  used.  As  soon  as  they 
have  been  taken  out  others  are  put  in. 

147.     LOSS  IN  DISTILLATION. 

The  loss  varies  from  10  to  25%  of  the  contained  zinc,  and 
occurs  in  several  ways.  The  residue  or  ashes  discharged  from  the 
retorts  retains  2  to  10%  zinc  in  the  form  of  undecomposed  zinc 
sulphide  and  zinc  oxide  that  escaped  reduction.  The  residue  is 
higher  in  zinc  at  the  extreme  front  of  the  retort.  The  lower  retorts 
in  a  direct-fired  furnace  are  hotter,,  and  hence  the  zinc  is  better 
expelled  from  them.  Thus  in  a  Belgian  furnace,  the  residue  from 
the  upper  row  contains  9.15%  zinc,  from  the  middle  row  4.67%, 
and  from  the  lower  row  but  2.28%.  On  the  other  hand  the  retorts 
of  the  topmost  row  last  90  days,  while  those  of  the  lowest  row 
last  only  6  days. 

Since  in  good  work  the  residue  contains  3  to  5%  zinc,  it  follows 
that  the  lower  the  grade  of  the  ore,  the  greater  is  the  percentage 
of  loss  figured  on  the  original  content.  Thus,  on  a  Joplin  ore  of 
60%  zinc,  we  expect  a  recovery  of  90%,  while  in  upper  Silesia, 
where  the  ore  carries  25%,  but  75%  is  recovered.  A  new  retort 
does  not  begin  to  give  the  maximum  output  of  zinc  until  it  has 


410  THE    METALLURGY 

been  in  use  several  days,  due  to  the  fact  that  the  clay  absorbs 
metal,  forming  zinc  aluminate,  an  artificial  zinc  spinel.  It  imparts 
a  deep  purplish  blue  color  to  the  retort,  and  an  old  retort  may 
contain  6%  or  more  of  zinc.  Zinc  is  lost  by  filtration  through  the 
walls  of  the  retorts  as  a  result  of  the  porosity  and  chimney-draft 
in  direct-fired  furnaces.  With  gas-fired  furnaces,  in  the  combustion 
chamber  there  is  a  little  pressure  which  is  one  of  the  advantages  of 
gas-firing.  Filtration  loss  is  greatly  diminished  by  using  retorts 
made  under  high  hydraulic  pressure,  and  by  the  addition  of  coke- 
dust  to  the  clay  mixture,  to  make  the  walls  of  the  retort  dense 
and  impermeable.  The  escape  of  zinc  vapor  through  a  cracked 
retort  is  an  important  cause  of  loss.  In  direct-fired  furnaces  the 
loss  is  indicated  by  the  cessation  of  the  small  flame  issuing  from 
the  condenser,  and  between  firings  by  the  white  smoke  of  zinc  oxide 
escaping  at  the  top  of  the  chimney.  In  a  gas-fired  furnace  in 
such  cases,  the  flame  at  the  condenser  is  greater,  and  becomes 
brownish  red  instead  of  the  usual  bright  bluish-green.  The  action 
of  corrosive  slags  on  retorts  results  in  making  small  holes  in  them, 
and  zinc  may  escape  in  this  way  before  the  defect  is  discovered. 
The  action  on  the  retorts  is  more  destructive  at  nights  when  most 
of  the  zinc  has  come  away,  and  the  firing,  especially  on  the  lowest 
retorts,  is  severe.  Retorts  weakened  by  the  action  of  the  corrosive 
slag  are  apt  to  develop  holes,  and  such  retorts  are  said  to  have  been 
'butchered.' 

When  the  condensers  are  removed,  at  the  end  of  the  distillation, 
there  is  a  loss  of  the  zinc-vapor  still  remaining,  which  escapes  and 
burns  to  zinc  oxide.  The  escape  of  zinc-vapor  during  the  retorting 
is  another  loss,  and  the  temperature  of  the  condenser  must  be  nicely 
regulated.  If  cold,  a  portion  of  blue-powder  results,  due  to  the 

.  condensation  of  the  zinc-vapor  in  powder  instead  of  in  liquid  form. 
If  hot,  then  zinc  escapes  condensation  and  burns  at  the  mouth  of 
the  condenser  with  the  greenish-white  flame  that  is  characteristic 
of  zinc.  Thus  the  regulation  of  the  temperature  of  the  condenser 
is  no  easy  matter.  The  eu^^^jL^  to  maintain  a  temperature  of 

jflf500°C.  or  fifctP  above  the  a3±»g7point  of  zinc.  Escaping  zinc 
may  be  partly  saved  by  the  use  of  prolongs  or  sheet-iron  cones 
that  fit  over  the  end  of  the  condenser.  There  is  less  advantage 
in  using  them  than  one  would  expect,  since  the  material  thus 
removed  is  in  the  form  of  blue-powder  that  must  be  re-distilled. 

148.     COST  OF  SMELTING. 

We    give    herewith    the    cost    in    Kansas    of    smelting    blende- 


OF    THE    COMMON    METALS.  411 

concentrate  per  ton  of  raw  ore.     The  roasted  ore  is  assumed  to  be 
86%  of  the  weight  of  the  raw  ore. 

COST  OF  SMELTING  PER  TON  OF  DRY  UNROASTED  CONCENTRATE. 


Coal. 

Natural 

gas. 

Hand 

Mech. 

Hand 

Mech. 

roast- 

roast- 

roast-     roast- 

ing. 

ing. 

ing. 

ing. 

Labor    (except   for   repairs  and    renewals) 

$6.62 

$5.82 

$4.70 

$4.25 

Fuel    (3   tons  at  75c  ) 

2.25 

2.25 

Reduction  material   ($1  per  ton)  

0.47 

0.47 

0.60 

0.60 

Clay    for    retorts,    condensers,    etc.,    at    0.1    ton, 

,  per  ton  of  ore  

0.26 

0.26 

0.26 

0.26 

Repairs,    renewals,    and    sundry    supplies     (in- 

cluding starting  furnace  in  operation).... 

0.75 

0.75 

0.75 

0.75 

$10.36          $9.56          $6.31          $5.86 

In  the  above  table  the  assumption  is  made  that  the  gas  (natural 
gas  being  used)  costs  nothing.  Since  we  have  to  meet  the  cost  of 
acquiring  the  land  for  natural  gas  and  that  of  pipe  line,  etc., 
additional  cost  for  gas  must  be  allowed. 

149.     THE  SADTLER  PROCESS. 

To  counteract  the  corrosive  action  of  the  bases  already  referred 
to,  Prof.  Benjamin  Sadtler  patented  a  process  that  consists  in  lining 
the  retort  with  basic  material,  preferably  chromite,  making  a  layer 
on  the  interior  of  the  retort  l/$,  in.  thick.  The  chromite  is  crushed 
to  pass  a  20-mesh  screen  and  the  dust  is  screened  out,  leaving  a 
coarse  granular  powder.  The  interior  of  the  retort  is  painted 
with  a  solution  of  water-glass,  and  while  wet,  sprinkled  with  the 
powder,  in  the  same  way  that  a  house  is  painted  and  sanded  to 
imitate  stone.  When  the  first  coat  is  dry  another  is  put  on,  the 
two  coats  giving  the  required  thickness.  The  further  drying  and 
annealing  is  done  as  in  making  regular  retorts.  The  lining  fuses 
to  the  interior  surface  of  the  retort,  and  doubles  the  life  (making 
80  days),  so  that  the  retort  fails  by  cracking,  or  by  gradually 
accumulating  residue  rather  than  by  corrosion.  The  retorts  are 
especially  adapted  to  treating  ore  containing  iron  and  lead  (such 
as  the  Leadville  ores).  The  residue,  still  retaining  the  precious 
metals  as  well  as  lead  and  the  excess  of  iron,  is  a  desirable  product 
for  silver-lead  smelting.  The  cost  for  putting  in  the  lining  is 
estimated  to  be  25c.  per  retort. 

The  retorts  have  been  used  at  the  works  of  the  Cherokee-Lanyon 
Spelter  Co.,  in  eastern  Kansas,  where  Colorado  ores,  containing 
iron  and  lead,  are  successfully  treated. 


PART  IX.     REFINING 


PART  IX.     REFINING. 

150.     PHENOMENA  UNDERLYING  THE   REFINING   OF 

METALS. 

It  is  found  in  analysis  that  the  separation  of  a  metal  from 
other  metals  or  from  contained  impurities  is  seldom  complete.  It  is 
difficult  and  commercially  impracticable  to  obtain  metals  entirely 
pure,  so  that  metals  that  come  on  the  market  still  contain  small 
amounts  of  impurity.  Metals  thus  prepared  are  graded  according 
to  quality,  and  command  prices  according  to  the  gr.ade.  Thus 
Lake  copper  commands  the  highest  price  of  copper  because  of  the 
purity  and  toughness,  while  casting-copper,  which  sells  at  ^c.  less 
per  pound,  is  used  for  making  brass. 

In  silver-lead  smelting  practice  the  slag,  no  matter  how 
thoroughly  settled  and  separated  from  the  matte,  still  contains  0.2 
to  0.3  oz.  silver  per  ton  and  0.3  to  0.4%  lead.  In  copper  refining, 
in  the  reverberatory  furnace  arsenic,  antimony,  and  bismuth,  occur- 
ing  in  the  crude  or  blister  copper,  are  retained  in  traces  after 
refining,  and  where  the  blister  copper  is  impure,  no  high-grade 
product  can  be  expected.  In  the  separation  and  deposition  of 
copper  by  electrolysis  at  low  current-density,  the  copper  is  of  high 
grade  even  though  impurities  are  in  solution  in  the  electrolyte ; 
nevertheless  traces  of  impurity  find  their  way  into  the  cathode 
copper,  though  to  less  extent  than  by  any  other  system  of  refining. 

In  the  refining  of  pig-iron  to  make  steel,  in  order  to  obtain 
satisfactory  quality,  impurities  must  be  removed  until  not  more  than 
0.10%  phosphorus  and  0.05%  sulphur  are  present,  otherwise  the 
steel  lacks  toughness  and  tenacity. 

151.     REFINING  LEAD  BULLION. 

Lead  containing  silver,  commonly  called  base-bullion,  is  refined 
by  the  Pattinson  or  by  the  Parkes  process.  Commonly  the  Parkes 
process  is  used.  The  object  in  either  process  is  to  effect  the. 
separation  of  the  silver  and  gold  from  the  lead. 

To  get  a  clear  idea  of  the  principles  of  refining  base-bullion 
(or.  work-lead  as  it  "is  called  in  Europe)  we  first  must  know  the 


416  THE    METALLURGY 

composition.      An    especially    base    quality    is    represented  by    the 
following  analysis : 

Per  cent. 

Lead    96.59 

Per  cent. 

Impurities :  Cu  0.82 

As  0.38 

Sb 0.71 

Fe  0.02 

S   0.14 

2.07 

Precious  metals:  Ag  (322  oz.  per  ton) 1.07 

Au  (0.20  oz.  per  ton) 0.0007 

—     1.07 

99.73 

It  is  seen  that  base-bullion  is  principally  lead.  The  problem  is 
to  soften  the  lead  by  removing  the  impurities,  and  then  to  separate 
the  gold  and  silver  from  the  purified  or  softened  lead.  In  studying 
the  process,  the  student  should  refer  to  Fig.  170. 

152.     SOFTENING. 

The  furnace. — Softening  is  performed  in  a  water-jacketed 
reverberatory-furnace,  Fig.  171.  The  rectangular  hearth  of  the 
furnace,  7  by  14  ft.  in  size,  is  surrounded  by  a  sheet-steel  double 
water-jacket,  shown  in  section  at  a  in  the  sectional  elevation,  c. 
The  jacket  resists  the  action  of  the  molten  litharge  formed  from  the 
lead  in  the  operation.  The  furnace  is  heated  by  the  fire-box,  having 
a  grate  4  by  5  ft.  in  dimensions,  so  that  a  high  temperature  can 
be  attained  in  the  furnace.  The  letters  c,  c  indicate  the  rear  working 
doors  and  &,  &;  &  the  front  doors  in  which  the  base-bullion  is  charged. 
At  the  front  of  the  furnace  the  tap-hole  e  is  provided,  through  which 
the  lead  is  tapped  when  the  charge  is  finished.  The  furnace 
communicates  to  a  stack  50  ft.  high  by  a  flue  at  the  front  end. 

Operation. — The  work  is  done  in  two  stages.  In  the  first  stage 
at  a  low  temperature,  the  copper  is  removed.  In  the  second,  at  a 
high  temperature,  the  arsenic  and  antimony  are  expelled,  after 
which  there  is  left  only  the  softened  lead  containing  the  precious 
metals. 

The  base-bullion,  in  charges  of  30  tons,  is  placed  in  the  furnace 
by  means  of  a  long-handled  paddle  or  'peel'  having  a  blade  2  ft. 
long  by  6  in.  wide.  The  bars  are  laid  one  at  a  time  upon  this 
and  placed  as  desired  in  the  furnace  being  piled  in  a  heap  on  the 
hearth.  The  doors  are  closed  and  the  bars  are  gradually  melted 
down,  the  dross  contained  in  the  bullion  rising  to  the  top.  When 
melted  the  heat  is  maintained  slightly  above  the  melting  point, 


OF    THE    COMMON    METALS. 


417 


but  not  higher.     In  about  two  hours  the  dross  that  has  risen  to 
the  top  is  carefully  skimmed,  by  means  of  a  long-handled  perforated 


skimmer,  and  removed  through  the  door  to  a  wheelbarrow  placed 
for  it.  The  dross,  residue,  or  skimming,  called  the  'copper  skim', 
consists  of  a  drossy  lead  containing  the  iron,  sulphur,  and 


418 


THE:  METALLURGY 


(especially  important)  most  of  the  copper  of  the  base-bullion.  The 
removal  of  these  completes  the  first  stage  of  the  process.  The 
liquated  dross  thus  skimmed,  which  may  amount  to  5%  of  the 
charge  or  1.5  tons,  consists  of  Pb,  62.4%  ;  Cu,  11.91%  ;  Ag,  0.17% 
(49  oz.  per  ton)  ;  As,  2.32%  ;  Sb,  0.98%  ;  Fe,  0.43%  ;  S,  4%  ;  and 
O,  1.87%.  Slag,  ash,  and  hearth  material  also  are  contained  and 
must  be  reckoned  in. 

The  heat  of  the  molten  bath  is  now  raised  to  a  bright  red ;  and 


Fig".    171.      SOFTENING   FURNACE. 


the  flame,  made  oxidizing  by  the  admission  of  an  excess  of  air 
through  the  thin  fire,  sweeps  over  the  surface.  Litharge  forms, 
and  the  antimony  and  arsenic  oxidize  and  enter  the  litharge  slag. 
The  litharge  at  this  temperature  has  a  corrosive  action  upon  the 
brick  lining,  hence  the  need  of  a  water-jacketed  furnace.  This 
stage  of  the  process  lasts  12  hours,  until  a  sample  of  the  lead 
taken  from  the  furnace  and  placed  in  a  mold  and  skimmed,  shows 
by  the  appearance  that  it  is  free  from  arsenic  and  antimony.  Before 


OF    THE    COMMON    METALS.  419 

the  antimony  is  removed  the  surface  of  the  molten  lead  will  'work,' 
or  show  oily  drops  moving  upon  it.  A  similar  phenomenon  is  seen 
in  the  first  stage  of  cupelling  base-bullion  high  in  antimony  and 
aisenic.  As  the  softening  proceeds  the  drops  become  fewer  and 
smaller,  and  finally  a  coating  is  seen  to  dull  the  surface  of  the  hot 
molten  lead,  indicating  the  completion  of  the  softening.  For  impure 
base-bullion  this  stage  is  of  more  than  12  hours'  duration,  and  the 
thick  layer  of  litharge  formed  retards  further  oxidation.  It  is 
best  then  to  draw  the  fire  and  to  cool  the  charge,  to  allow  the 
litharge  slag  on  the  top  to  solidify  above  the  liquid  lead  beneath. 
The  slag  is  skimmed  with  a  long-handled  perforated  skimmer 
(compare  with  Fig.  175),  and  the  charge  is  fired  again  if  necessary 
until  the  impurities  are  removed.  The  'antimony  skim'  consists  of 
the  antimonate  and  arsenate  of  the  lead  with  a  large  proportion  of 
litharge.  It  is  in  fact  an  impure  litharge. 

The  softened  lead  to  be  treated  either  by  the  Pattinson  or  the 
Parkes  process  for  the  removal  of  the  contained  gold  and  silver, 
is  now  tapped  into  the  desilverizing  kettle,  8  ft.  diam.  and  capable 
of  holding  30  tons  of  lead. 

153.     THE  PATTINSON  PROCESS. 

When  a  kettle  containing  molten  lead  is  allowed  to  cool  slowly, 
as  it  approaches  solidification,  crystals  of  lead  low  in  silver  separate. 
The  metal  that  remains  liquid  contains  the  larger  part  of  the 
precious  metal.  The  crystals  are  removed  with  a  perforated  ladle, 
melted  in  another  kettle,  and  allowed  to  cool.  Once  more,  crystals 
separate  that  are  low  in  silver,  the  mother  liquor  becoming  high 
in  silver.  If  the  liquid  portion  first  mentioned  be  transferred  to 
a  kettle  and  likewise  heated  and  then  allowed  to  cool,  the  same 
segregation  of  the  silver  into  the  liquid  part  continues.  We  can 
accordingly  arrange  a  series  of  kettles  containing,  at  one  end, 
low-grade  lead,  and  at  the  other  high-grade,  all  from  one  product. 
A  series  of  this  kind,  as  illustrated  by  practice  at  Eureka,  Nevada, 
gave  the  assays  quoted  in  the  table  below. 

Another  crystallization  would  reduce  the  silver  of  the  market 
lead  to  half  the  value  given.  The  rich  lead  could  be  directly 
cupelled  in  an  English  cupelling  furnace,  or  better,  treated  by  the 
Parkes  process  to  get  rich  silver-zinc  crust  for  retorting  and 
cupelling.  The  process  has  been  modified  -recently  by  Tredennick 
who  raises  each  kettle  by  hydraulic  power  above  the  adjoining  one  so 
that  the  mother  liquor  drains  from  one  kettle  to  the  next  through 


420  THE    METALLURGY 

a  strainer,  the  lead  being  cooled  near  the  solidification  temperature 
by  introducing  steam  upon  the  surface.  The  cost  of  operating  has 
been  greatly  decreased  in  this  way.  The  chief  advantage  of  the 
Pattinson  process  over  the  Parkes  is  that  it  gives  a  product  free 
from  bismuth.  In  the  Parkes  process  the  bismuth  follows  the 
lead.  Bismuth  is  injurious  in  lead  that  is  to  be  corroded  to  make 
white  lead,  and  it  may  be  necessary  to  employ  the  Pattinson  process 
for  making  a  corroding  lead  from  bismuth-bearing  ores. 

Market  Lead 
Kettle  oz.  Ag  per  ton. 

No.     1  1.25 

No.     2  2.5 

No.     3 5.0 

No.     4  9.0 

No.     5  .* 18.0 

No.     6  30.0 

No.     7  50.0 

No.     8  75.0 

No.     9 100.0 

No.  10  150.0 

*No.  11  450.0 

*Rich  Lead. 

154.     THE  PARKES  PROCESS. 

Operation. — The  softened  lead  from  the  softening  furnace  is 
tapped  into  a  hemispherical  cast-iron  kettle,  shown  in  Fig.  172 
and  173,  which  holds  30  tons  or  more  of  lead  or  the  full  charge 
from  the  softener.  The  kettle  is  set  in  brick-work,  and  is  heated 
from  a  fire-box  below.  In  modern  practice  kettles  are  made  large 
and  are  10  ft.  diam.  and  2  ft.  10  in.  deep,  holding  60  to  65  tons. 

The  principal  of  the  separation  of  silver  from  lead  depends  on 
the  affinity  of  silver  for  zinc  which  is  greater  than  for  lead.  Upon 
adding  and  thoroughly  mixing  in  a  small  amount  of  zinc  (\%  &£ 
4ho  whole),  the  'nna  takes  up  most  of  the  silver.  Zinc  has  a  greater 
affinity  than  lead  not  only  for  silver  but  for  gold  and  copper.  When 
the  molten  bath  is  allowed  to  stand  a  while,  the  zinc  being  lighter 
separates  and  rises  to  the  surface.  At  a  temperature  below  the 
melting  point  of  zinc  but  above  that  of  lead,  a  crust  forms  that 
can  be  skimmed  off.  Thus  the  silver  is  concentrated  in  a  small 
bulk  of  metal,  and  is  later  separated  from  the  rich  metal  by 
further  treatment. 


OF    THE    COMMON    METALS. 


421 


The   lead  from  the   softening  furnace   flows   along   a   cast-iron 

Another  crystallisation  could  reduce  tho  GJlvor  of  the  markci 
trough  to  the  kettle.  In  so  doing  a  litharge  dross,  called  'kettle 
dross'  forms,  and  collects  on  the  surface  of  the  metal,  and  is 
skimmed  off. 

The  molten  bath  is  next  heated  to  an  incipient  red  heat,  well 
above  the  melting  point  of  zinc  and  cakes  or  ingots  of  spelter  equal 
to  about  1.2%  of  the  weight,  or  720  Ib.  are  added.  The  quantity 
required  varies  with  the  richness  of  the  base-bullion  in  silver.  The 
added  zinc  quickly  melts. 

Desilverizing  machinery  is  now  much  used.  The  most  approved 
machine  is  the  Howard,  used  both  for  mixing  and  skimming.  Fig. 


Fig.   172.      HOWARD  MIXER. 

172  represents,  in  section  and  elevation,  the  kettle  and  the  apparatus 
used  for  intimately  mixing  the  molten  zinc  with  the  lead.  The 
machine  is  brought  to  the  kettle  by  an  overhead  crawl  h  and  is 
lowered  into  it  by  a  chain-block  hoist.  When  lowered  into  position 
(shown  in  section  Fig.  172)  the  screw  propeller  b  is  set  in  motion 
by  a  steam-driven  mechanism  to  produce  a  downward  flow  of 
molten  lead  in  the  sheet-iron  cylinder  a.  The  cylinder  has  neither 
top  nor  bottom,  and  being  submerged  in  the  lead,  a  circulation  is 
started,  the  lead  flowing  in  over  the  top  of  the  cylinder.  Thus  a 


422 


THE    METALLURGY 


thorough  mixing  of  the  content  of  the  kettle  is  assured.  In  a  few 
minutes  the  engine  is  reversed,  and  the  flow  is  made  upward  over 
the  edge  of  the  cylinder,  then  downward  to  the  bottom.  ^The  mixing 


Fig.    173.      HOWARD 
PRESS. 


Fig.    174.      HOWARD   PRESS 
(SECTION). 


is  continued  about  11  minutes,  after  which  time  the  stirring 
apparatus  is  bodily  hoisted  and  moved  to  one  side.  Several  kettles 
can  be  thus  served  by  one  mixer. 

The  content  of  the  kettle  is  now  allowed  to  cool  two  hours  or 


OF    THE    COMMON    METALS.  423 

more.  The  light  zinc  rises  to  the  top  and  carries  the  silver,  gold, 
and  copper  with  it.  Finally  when  the  temperature  falls  below  the 
melting  point,  a  half-fused,  mushy  crust  or  layer  forms  upon  the 
lead.  The  crust  consists  of  65%  Pb,  10%  Ag  and  Au,  3%  Cu, 
and  22  to  24%  Zn. 

Fig.  173  is  an  elevation  of  the  Howard  press  by  which  the  zinc 
crust  is  removed  from  the  lead.  Fig.  174,  another  elevation,  shows 
also  a  section  of  the  cast-iron  pot  into  which  the  press  is  about 
to  be  lowered.  In  principle  the  machine  is  like  a  cheese-press.  The 
apparatus  is  lowered  into  the  lead  until  the  top  edge  of  the  cylinder 
a  is  but  slightly  above  the  surface.  The  plunger  or  follower  c  is 
raised,  and  the  zinc-crust,  as  it  is  skimmed  from  the  surface,  is 
put  in  it  by  means  of  the  perforated  skimmer,  Fig.  175. 


Fig.   175.      SKIMMER. 

The  press  thus  expedites  the  skimming.  While  one  man  is 
skimming  and  putting  the  skimming  into  the  press,  another  assists 
by  pushing  the  crust  to  one  place  with  a  wooden  rabble.  When 
full,  the  press  is  raised  and  the  surplus  lead  begins  to  run  out  of 
the  half-inch  holes  in  the  hinged  bottom  1).  The  plunger  c  is  brought 
down,  squeezing  out  more  of  the  lead,  and  leaves  the  remaining, 
mushy  half-fluid  mass  nearly  free  from  lead,  of  the  composition 
given  above.  The  press  is  now  run  to  one  side  over  a  floor  paved 
with  cast-iron  plates,  the  hinged  bottom  &  is  dropped  by  releasing 
the  catch,  and  the  zinc-crust  is  pushed  out  by  continuing  the  down- 
ward movement  of  the  plunger.  The  crust  falls  upon  the  cast-iron 
floor-plate,  and  while  soft  is  readily  broken  with  hammers  into 
lumps  the  size  of  the  fist.  Meanwhile  the  hinged  bottom  is  closed, 
and  the  press  returned  to  the  kettle  and  is  opened  to  receive  more 
skimming.  These  operations  continue  until  the  surface  of  the  lead 
is  well  skimmed.  The  crust  amounts  to  3000  Ib.  and  contains  90%  of 
the  silver  originally  in  the  softened  base-bullion,  resulting  in  a 
concentration  of  twenty  into  one. 

This  first  'zincking'  removes  all  the  gold  and  copper  for  which 
zinc  has  a  great  affinity.  It  does  not  remove  all  the  silver,  and 
the  operation  must  be  repeated  once  or  twice  more  before  the  silver 


424 


THE    METALLURGY 


content  is  diminished  to  the  fraction  of  an  ounce  per  ton  beyond 
which  it  does  not  pay  to  go.  Of  the  1.8%  zinc  needed,  first  is  added 
%  of  the  zinc  or  1.2%,  then  ~y±,  or  0.45%,  and  finally  the  remaining 
Vi2>  or  0.15%  ;  or  900,  270,  and  90  Ib.  respectively. 

The  desilverized  lead  remaining  in  the  kettle  after  the  last 
skimming  retains  0.6  to  0.7%  zinc  and  traces  of  arsenic  and 
antimony,  all  of  which  must  be  removed  before  the  lead  is  suitable 
for  market.  This  is  done  by  siphoning  or  tapping  the  metal  from 
the  kettle  into  a  reverberatory  furnace  similar  in  construction  to 
the  softening  furnace.  Here  the  charge  is  brought  up  to  a  bright- 


176.      MOLDING  MARKET-LEAD. 


red  heat,  the  zinc  is  volatilized  and  burned  off,  and  litharge  forms 
as  a  slag  upon  the  surface  of  the  lead.  The  operation  takes  six 
hours,  and  the  completion  when  the  zinc  has  been  expelled,  is 
known  by  taking  a  sample  of  lead  in  a  mold  and  observing  the 
appearance  of  the  surface  as  the  metal  solidifies.  The  furnace  is 
allowed  to  cool  until  the  litharge-slag  is  solid  and  can  be  skimmed. 
Finally  the  lead  is  tapped  into  a  market-kettle  similar  to  the 
desilverizing  kettle.  This  is  the  reservoir  from  which  it  is  drawn 
to  be  cast  into  molds.  The  molds,  50  in  number,  standing  in  a  semi- 
circle as  shown  in  Fig.  176,  hold  100  Ib.  of  lead  each,  and  are  con- 
veniently mounted  on  two  wheels  by  which,  when  full  and  cool, 
they  are  transferred  to  the  adjoining  floor.  There  the  lead  is  tilted 


OF    THE    COMMON    METALS. 


425 


out  and  the  molds  at  once  returned  to  the  semi-circle  to  be  used 
again.  The  lead  is  withdrawn  from  the  kettle  by  means  of  a  siphon. 
It  descends  into  a  small  cast-iron  pot,  into  which  is  screwed  the 
21/2-in.  pipe  that  delivers  it  to  the  50  molds,  the  pipe  being  quickly 
moved  from  mold  to  mold  as  filled,  without  interrupting  the  flow. 


(b)    ' 


FABER   DU  FAUR  RETORT. 


At  the  end  of  each  round  the  flow  is  interrupted  only  to  carry 
back  the  end  of  the  pipe  to  the  first  mold  of  the  series,  which  mean- 
while has  been  emptied  and  replaced.  The  100-lb.  pigs,  or  bars,  are 
the  desilverized  lead  of  commerce. 

Dry  steam  may  be  blown  into  the  molten  lead  in  the  kettle  to 
refine  it.  It  is  introduced  by  means  of  a  pipe  inserted  deep  beneath 
the  surface.  The  constant  agitation  produced  by  the  steam  brings 


426 


THE    METALLURGY 


the  metal  in  contact  with  the  air  and  oxidizes  it  and  the  remaining 
impurity.  It  is  softer  than  ordinary  desilverized  lead,  and  is  easily 
corroded  by  the  acetic  acid  used  in  making  white  lead.  It  is  accord- 
ingly called  'corroding  lead.' 


Fig.   1' 


ENGLISH   CUPELLING   FURNACE. 


Section   C.D. 


V3 


=t 


— 0 


Fig.    179.      ENGLISH  CUPELLING   FURNACE    (DETAILED   SECTIONS). 

Retorting. — Referring  to  the  flow-sheet  (Fig.  170),  we  see  what 
becomes  of  the  crust  or  skimming  that  results  from  the  first  zinck- 
ing.  The  material  is  in  lumps,  containing  22  to  24%  zinc.  It  is 
charged  with  charcoal  breeze,  into  bottle-shaped  retorts  /,  Fig.  177, 


OF   THE    COMMON   METALS.  427 

each  holding  1200  Ib.  zinc-crust.  The  figure  represents  at  a  a 
sectional  elevation  through  the  retort,  at  b  a  transverse  section,  and 
at  c  an  elevation  of  a  Faber  du  Faur  tilting  retort-furnace.  The 
retort  rests  upon  a  narrow  arch,  and  carries  a  grate  upon  which 
rests  a  coke  fire  that  fills  the  furnace  and  covers  the  retort  /.  The 
products  of  combustion  escape  by  an  outlet-port  at  the  back  to  a 
stack  of  good  draft.  The  coke  is  fed  through  a  hole  in  the  roof 
of  the  furnace,  and  is  poked  down,  and  kept  in  vigorous  combustion, 
so  that  a  yellow  heat  (1000°C.)  is  attained.  A  condenser,  made  by 
cutting  off  the  end  of  an  old  retort  (as  shown  at  c  Fig.  177),  collects 
the  zinc  vapors  distilling  from  the  charge,  the  condensed  zinc  being 
drawn  into  molds  x  through  a  one-inch  hole,  bored  through  the 
bottom  edge  of  the  condenser.  When  distillation  is  complete,  the 
condenser  and  the  supporting  truck  is  removed,  the  furnace  is  tilted 
or  revolved  by  means  of  a  lever  on  the  trunnions,  and  the  remaining 
'rich-lead'  is  poured  into  molds  like  those  used  for  molding  market- 
lead.  The  rich  lead  still  retains  zinc,  copper,  and  impurities,  taken 
into  the  crust  at  the  first  zincking. 

Cupelling. — To  obtain  the  silver  (and  gold)  from  the  alloy,  the 
English  cupelling-furnace,  Fig.  178,  179,  is  used.  The  principle  of 
the  action  is  much  like  that  of  cupellation  in  assaying,  except  the 
litharge  here  not  only  saturates  the  cupel,  but  flows  from  it  as 
fast  as  formed.  Fig.  179  is  a  sectional  elevation  of  the  furnace, 
showing  the  fire-box  a  where  a  long-flaming  coal  is  burned,  the 
products  of  combustion  passing  to  the  chimney  by  the  port  ~b  and 
an  underground  flue.  The  flame  plays  over  the  hearth  called  a 
'test,'  a  large  cupel,  not  shown  in  Fig.  179,  but  shown  mounted 
on  the  carriage  in  Fig.  178.  The  test  is  lowered  by  the  jack  screws, 
removed  on  the  carriage,  and  another  is  put  in,  when  the  first  is 
consumed.  The  test  is  hollowed  like  a  cupel,  to  hold  a  shallow  bath 
of  molten  lead  3  in.  deep.  .Fig.  180  shows  two  views  of  the  test  and 
the  supporting  truck  including  a  view  of  an  inverted  truck  and  test. 
In  the  elevation  of  the  furnace,  Fig.  178,  is  seen  the  overhead  pipe, 
branching  to  the  ash-pit,  to  supply  undergrate-blast,  and  to  an 
opening  at  the  back  of  the  furnace  where  a  tuyere  is  inserted,  by 
which  a  stream  of  air  is  brought  to  play  upon  the  surface  of  the 
molten  red-hot  bath  of  rich  lead.  The  air  oxidizes  the  lead  to 
litharge.  Other  impurities  are  oxidized  and  enter  the  litharge-slag 
and  are  carried  away  with  it.  The  molten  litharge,  as  it  forms,  runs 
along  a  shallow  groove  or  channel  in  the  top  of  the  front  edge 
of  the  cupel  or  test.  A  door  at  the  front  can  be  lifted  to  inspect 
the  oneration,  or  to  cut  the  channel  as  needed.  At  the  rear  are 


428 


THE    METALLURGY 


provided  two  ports  of  a  size  to  permit  inserting  two  bars  of  rich 
lead  that  are  pushed  in  as  fast  as  the  cupellation  proceeds.  The 
ends  of  the  bars  melt  and  supply  the  lead.  The  litharge  stream  is 
the  size  of  a  lead  pencil,  and  falls  into  a  small  slag-pot  beneath.  The 
lead  is  fed  at  the  rate  of  one  to  two  tons  daily  until  the  bath  has 
become  rich  in  silver,  when  the  feeding  of  the  lead  must  be  stopped. 
Oxidation  then  is  continued,  cutting  the  channel  deep  to  allow  the 
remaining  litharge  to  flow  out,  and  finally  the  mirror-like  bath  of 
silver  appears.  The  fire  must  keep  the  temperature  above  the  melt- 
ing point  of  the  silver.  At  the  last,  a  shovelful  of  bone-ash  is  thrown 
on  the  bath  to  absorb  the  remaining  trace  of  litharge  as  it  fo^ms 
on  the  surface.  This  is  skimmed,  and  the  silver  is  then  ready  to  be 
ladled  out  or  tapped,  commonly  into  cast-iron  molds  each  holding 


sft 

i 


Fig.    180.      TEST   FOR  ENGLISH   CUPELLING  FURNACE. 

1000  oz.  silver.  The  silver  is  then  subjected  to  the  acid-parting 
operation  to  be  described  later. 

The  copper-skimming,  which  is  the  first  obtained  from  the 
softening  furnace,  is  returned  to  the  blast-furnace  where  the  sulphur 
of  the  charge  combines  with  the  copper  and  removes  it  as  matte. 
The  rest  of  the  skimming,  mostly  lead,  containing  silver  and  gold, 
is  reduced  to  base-bullion.  The  third  skimming  of  the  softening 
furnace,  if  any,  is  returned  to  the  blast-furnace  since  it  contains 
but  little  antimony. 

The  second  softening  skimming  or  antimony-skim,  containing 
15  to  25%  antimony,  goes  to  a  small  reverberatory  furnace,  8  by 
12  ft.  hearth  dimensions  and  10  in.  deep,  built  like  a  softening 
furnace  and  called  a  precipitating  furnace.  Here  it  is  melted,  with 
a  reducing  flame  into  a  slag.  Charcoal  is  added,  and  stirred,  to 
reduce  or  precipitate  part  of  the  lead  of  the  slag.  The  lead,  falling 


OF    THE    COMMON    METALS.  429 

to  the  bottom  of  the  bath,  carries  down  the  silver  of  the  slag.  When 
the  reaction  is  complete  the  supernatant  slag  is  tapped  into  slag- 
pots,  and  the  lead  is  tapped  into  a  kettle  at  a  lower  level  and  molded 
into  bars.  This  precipitated  lead-bullion  is  returned  to  the  softening 
furnace  to  be  softened  and  desilverized.  The  antimonial  slag, 
containing  about  6  oz.  silver  per  ton,  when  accumulated,  is  smelted 
in  a  small  blast-furnace  to  reduce  it  to  antimonial  lead  of  20%  Sb, 
which  is  sold  to  the  type  founders.  The  slag  is  rejected. 

155.     VARIATIONS  IN  METHODS  OF  REFINING  BASE- 
BULLION. 

Instead  of  removing  the  gold  and  silver  together,  in  the  first 
zinc-crust,  it  is  possible  to  take  the  gold,  copper,  and  a  part  of  the 
silver  by  a  separate  zincking,  in  which  a  moderate  amount  of  zinc 
(250  to  350  Ib.)  is  used.  The  separation  is  due  to  the  fact  that 
zinc  has  a  greater  affinity  for  gold  and  copper  than  for  silver.  We 
thus  obtain  a  zinc-crust  containing  all  the  copper,  gold,  and  some  of 
the  silver.  From  this  crust,  upon  retorting  and  cupelling,  a  dore 
bar  (a  silver  bar  containing  the  gold)  is  produced.  The  second 
zincking  accordingly  needs  but  600  Ib.  zinc,  and  removes  most  of  the 
remaining  silver.  A  final  zincking  makes  the  lead  practically  clean. 

The  second  zinc-crust  yields  a  silver  bar  free  from  gold,  on  which 
the  expense  of  parting  is  saved,  and  the  full  price  of  commercial 
bar-silver  is  obtained. 

In  the  practice  of  early  days,  on  clean  base-bullion,  made  from 
clean  ores,  softening  was  performed  in  the  kettle.  The  bars  were 
melted  at  a  low  temperature,  and  the  dross  was  made  as  dry  (free 
from  lead)  as  possible  before  skimming.  The  skimmed  lead  was 
brought  to  a  red  heat  by  vigorous  firing,  and  stirred  to  furnish 
contact  with  the  air.  Arsenic  and  antimony  were  present  in  limited 
quantity  and  could  be  removed  in  this  way.  For  ordinary  base- 
bullion  the  method  is  inadequate,  and  the  softening-furnace  must 
be  used. 

When  lead  has  been  desilverized,  the  contained  zinc  can  be 
removed  in  the  desilverizing  kettle  without  the  use  of  a 
reverberatory  calcining  furnace.  This  is  accomplished  by  heating 
the  lead  in  the  kettle  to  a  high  temperature  and  rabbling  by  means 
of  dry  steam  or  compressed  air,  admitted  into  the  kettle  through  a 
pipe  that  leads  to  the  bottom.  As  the  steam  rises  it  agitates  the 
hot  metal  and  brings  it  in  contact  with  the  air  above  the  kettle. 
The  zinc  is  thus  in  part  volatilized,  in  part  oxidized  with  a  little 


430  THE    METALLURGY 

of  the  lead.     A  powdery  litharge  is  formed  that  is  skimmed  off, 
leaving  market  lead. 

In  the  old  way  of  stirring  the  zinc  into  the  charge  of  the 
desilverizing  kettle,  the  work  was  done  by  means  of  paddles  (shown 
in  Fig.  181)  having  a  handle  7  ft.  long.  Two  men  worked  with  the 
paddles,  one  on  each  side  of  the  kettle.  The  paddle  blades  were 


Fig.    181.      STIRRING  PADDLE. 

moved  as  in  the  operation  of  rowing,  thoroughly  mixing  the  zinc 
into  the  lead  in  15  minutes  constant  stirring. 

Where  the  Howard  press  was  not  used,  the  first  or  silver-crust, 
from  the  desilverizing  kettle,  was  transferred  to  another  kettle  and 
kept  at  a  low-red  heat.  In  consequence  there  was  a  separation  of 
a  liquid  portion  that  was  retained  and  sent  to  the  next  charge,  the 
dry  residue  being  the  copper  skimming.  The  bars  were  put  into  the 
softening  furnace  to  complete  the  refining. 

156.     COST  OF  REFINING  BASE-BULLION. 

The  actuual  cost  of  refining  base-bullion  is  as  follows : 
Prime    or    flat-cost,    of    softening 

and  refining   $5.00  to     $6.00 

General  expense    3.00  to       3.00 

Loss     in     metals     and-    incidental 

expenses   1.70  to       3.00 


Total    $9.70  to  $12.00 

157.     COPPER  REFINING. 

Blister  copper,  or  black  copper,  whether  produced  in  the  blast- 
furnace, the  converter,  or  the  reverberatory  furnace,  or  by  melting 
the  concentrate  from  the  native  copper  ore  of  the  Lake  Superior 
region,  still  contains  impurity,  principally  arsenic  with  sulphur 
and  iron,  and  all  impurity  must  be  removed  by  refining.  If  the 
copper  contains  gold  and  silver  in  quantity  to  warrant  (20  to  40  oz. 
silver  per  ton),  it  is  melted  without  attempting  to  refine,  and  cast 
into  anodes  that  then  are  subjected  to  electrolytic  refining.  If 
there  is  but  little  precious  metal  in  the  copper  it  may  be  directly 
refined  in  the  copper  refining-furnace.  Fig.  182  is  a  sectional 


OF    THE    COMMON    METALS. 


431 


3levation,  and  Fig.  183  a  sectional  plan  of  a  40,000  to  50,000  Ib. 
copper-refining  furnace.  It  is  14  by  19  ft.  hearth  dimensions,  and 
has  a  fire-box  5%  by  6^2  ft.,  or  30  ft.  area,  and  carries  a  fire-bed 

L 


4%  ft.  thick.  The  hearth,  2V2  ft.  deep,  has  a  brick  or  a  sand  bottom. 
If  of  sand,  the  bottom  is  carefully  smelted  in.  Beneath,  the  hearth 
is  vaulted  for  ventiliation.  The  bridge,  5  ft.  wide,  is  strengthened 
by  a  double  conker-plate,  and  on  either  side  and  in  the  roof  over 
the  fire  bridge,  are  ports  that  are  opened  when  an  oxidizing  flame 


432  THE    METALLURGY 

is  desired.  In  the  elevation,  at  the  front  end,  is  to  be  seen  the 
outlet-flue  that  leads  to  the  stack  or  chimney.  The  chimney  is 
close  to  the  furnace  but  is  not  shown  in  the  plan.  The  charge  of 
ingots  of  blister-copper  is  put  in  at  the  side  door.  The  door  is 
then  tightly  closed,  and  vigorous  firing  follows.  The  charge  melts 
after  several  hours.  The  front  door  is  then  opened,  and  whatever 
slag  formed  during  the  melting  is  skimmed. 

Next  follows  the  rabbling,  the  object  of  which  is  to  oxidize  a 
portion  of  the  copper  and  the  impurities  with  it.  The  operation 
years  ago  consisted  in  striking  the  surface  of  the  bath  with  a  rabble 
in  such  a  way  as  to  splash  the  metal  and  agitate  it,  thus  exposing 
it  to  the  action  of  the  air.  The  present  way  is  to  insert  a  %-in.  pipe 
just  beneath  the  surface  of  the  metal  and  force  compressed  air 
through  it  to  agitate,  and  at  the  same  time  to  oxidize  it.  The  air-ports 
of  the  furnace  also  are  opened  and  the  flame  is  made  an  oxidizing 
one.  The  action  proceeds  to  the  stage  of  'set  copper,'  Cu2O  having 
been  by  this  time  formed,  and  in  part  dissolved  in  the  copper.  Iron, 
sulphur,  and  arsenic  partly  volatilize,  and  partly  oxidize  and  enter 
the  slag  that  is  formed  at  the  same  time,  this  is  skimmed  off. 

The  copper  oxide  must  be  removed  by  poling.  This  is  a  reducing 
action  in  which  the  air-ports  are  closed  to  give  a  reducing  flame,  and 
spruce  or  poplar  poles  are  inserted  at  the  front  door  into  the  metal. 
The  outer  end  of  the  pole  is  raised  to  force  the  butt-end  beneath 
the  surface  of  the  metal.  At  the  same  time  a  wheelbarrow-load  of 
charcoal  is  thrown  in  to  cover  the  surface,  to  exclude  air,  and  to 
reduce  cuprous  oxide.  As  the  hydrocarbon  of  the  wood  is  evolved 
and  the  moisture  evaporates,  that  is  as  the  wood  burns,  reduction 
takes  place.  The  operation  requires  an  hour  or  two.  Additional 
poles  are  inserted  to  replace  those  consumed.  Samples  of  a  few 
ounces  of  the  copper  are  removed  in  a  small  ladle  from  time  to 
time  and  examined  to  note  the  progress  of  reduction.  The  'tough 
pitch,'  (the  point  at  which  the  cuprous  oxide  is  completely  reduced 
to  metallic  copper)  is  the  end  in  view. 

The  charge  is  now  ready  for  dipping  or  ladling.  Hand-ladles, 
holding  25  Ib.  or,  large  'bull  ladles,'  holding  200  Ib.  and  carried 
by  an  overhead  trolley  or  crawl,  are  used.  The  dipping  or  molding 
consumes  three  hours.  The  copper  is  kept  hot  by  occasional  firing, 
and  by  keeping  the  surface  of  the  metal  covered  with  charcoal. 
The  charcoal  serves  also  to  keep  the  copper  in  pitch,  or  in  the 
condition  of  tough  copper.  The  molds  into  which  the  copper  is 
poured  from  the  ladles  are  of  the  shape  required  by  the  trade. 
There  are  required;  ingots  or  bars  suited  to  re-melting  for  making 


OF    THE    COMMON    METALS.  433 

brass ;  wire  bars  of  a  form  convenient  for  roiling  into  wire ;  and 
rectangular  cakes,  often  18  in.  square  and  4  in.  thick,  but  also  of 
dimensions  giving  2000  to  4000  Ib.  weight.  The  size  of  the  cake 
is  suited  to  the  size  of  the  sheets  of  copper  into  which  they  are  to 
be  rolled. 

158.     MELTING  AND   REFINING  LAKE   SUPERIOR  COPPER. 

The  product  from  which  copper  is  made,  in  the  Lake  Superior 
region,  is  a  concentrate  (called  locally  'mineral'),  which  averages 
70%  copper  in  native  form,  accompanied  with  a  self-fluxing  or 
fusible  gangue.  In  addition  there  occur  pieces  of  copper  of  different 
sizes,  from  that  of  the  fist  to  several  tons  in  weight,  called  mass- 
copper.  The  small  pieces  are  handled  easily,  and  are  shipped  to  the 
smelting  works  in  barrels.  It  is  called  barrel-work.  The  larger 
pieces  called  'mass  copper'  or  simply  'mass,'  are  70%  copper.  For 
the  large  pieces  that  cannot  be  charged  at  the  side  doors,  a  hatch- 
opening  with  a  clamped  brick  cover  is  provided.  Large  pieces  are 
raised  by  a  crane  and  charged  through  the  hatch,  also  concentrate 
or  mineral  to  make  a  charge  of  36,000  to  40,000  Ib.  The  charging 
takes  place  immediately  after  the  dipping  and  the  repairing  or 
fettling  the  furnace. 

The  furnace  is  now  closed,  and  firing  proceeds  for  several  hours. 
As  the  charge  melts  and  slag  forms,  it  is  skimmed  until  the  metal 
is  completely  melted  and  the  surface  is  clear.  The  operation  of 
refining  then  continues  as  has  been  described  above.  The  slag 
contains  15  to  25%  copper,  partly  as  entrained  prills  and  flakes,  and 
partly  as  cuprous  oxide.  This  slag  is  smelted  in  a  blast-furnace  with 
added  limestone  to  make  the  resulting  slag  fusible,  using  anthracite 
coal  and  a  portion  of  coke  for  fuel.  The  charge  consists  of  slag  2000 
Ib.,  limestone  600  Ib.,  and  anthracite  coal  400  Ib.,  and  the  slag 
resulting  contains  41%  SiO,,  25%  CaO,  and  22%  FeO.  It  is  the 
practice  at  one  works  to  add  much  small  mass-copper  and  barrel- 
work  to  the  charge,  the  idea'  being  that  as  the  copper  melts  and 
sinks  into  the  crucible  of  the  furnace  in  the  form  of  drops,  it  carries 
down  particles  of  reduced  copper  with  it.  The  products  of  the 
furnace  are  slag,  of  less  than  1%  copper,  and  cupola  blocks  (impure 
blister-copper) . 

In  recent  practice  in  the  Lake  Superior  country  the  operation  of 
melting  is  performed  in  one  furnace  and  refining  in  another.  A 
furnace,  18  by  40-ft.  hearth  area  melts  100  tons  of  mineral  of  67% 
copper  in  24  hours  using  30  tons  of  coal.  The  charge  is  supplied  in 
two  portions,  the  second  following  as  soon  as  the  first  is  melted. 


434  THE    METALLURGY 

Slag  is  removed  several  times  during  the  period,  and  the  copper 
when  free  from  slag  is  tapped  into  a  refining  furnace  at  a  lower 
level.  The  refining  is  done  in  a  14  by  22-ft.  furnace  carrying  a  deep 
charge  of  copper.  The  charge  is  made  from  copper  scrap,  high- 
grade  'mineral'  of  over  80%  copper,  and  mass-copper.  It  is  melted, 
and  when  ready  the  charge  from  the  melting  furnace  is  added,  so 
that  the  total  charge  consists  of  275,000  to  300,000  Ib.  copper.  The 
large  quantity  of  copper  is  refined  by  rabbling  or  poling.  To  cast 
it  in  a  reasonable  time,  a  casting  machine  is  employed  which  trans- 
forms the  copper  into  commercial  shapes  at  the  rate  of  50,000  Ib. 
per  hour. 

The  cupola-blocks,  referred  to  above,  wThich  constitute  the 
product  of  the  blast-furnace  smelting  of  the  slag  from  the  rever- 
beratory  refining-furnace  are  melted  and  refined  to  produce  a  low- 
grade  copper  called  'casting-copper.'  The  cupola-blocks  are  impure 
and  contain  so  much  arsenic  that  it  is  practically  impossible  to 
remove  it  all.  Other  impurities  (iron  and  sulphur)  are  eliminated. 

159.     ELECTROLYTIC  COPPER  REFINING. 

Blister-copper  from  reduction  works  in  the  Western  States  is 
profitably  treated  by  electrolytic  refining.  Not  only  can  a  pure 
copper  be  produced  from  impure  material,  but  the  gold  and  silver 
in  the  blister-copper  can  be  separated  and  recovered.  The  copper 
from  the  reduction  works,  in  the  form  of  rough  ingots,  is  sampled 
to  determine  the  content  of  precious  metal  and  the  copper.  The 
electrolytic  refining  of  copper  is  well  illustrated  in  the  practice 
of  the  Raritan  Copper  Co.,  Perth  Amboy,  N.  J.  Sampling  is  done 
by  drilling  each  ingot  or  bar  with  a  %-in.  drill,  the  drillings  being 
then  mixed  for  the  sample  of  the  lot. 

The  copper  after  being  sampled  is  sent  to  the  re-melting-furnace 
which  holds  150,000  Ib.  copper.  It  is  melted,  poled,  and  cast  into 
smooth  anodes  by  means  of  a  Walker  casting-machine.  The  anodes, 
weighing  300  Ib.  each,  coming  from  the  machine,  are  trimmed  to 
remove  fins  or  projections  of  copper,  and  are  loaded  on  trucks 
holding  22  anodes  in  a  rack. 

The  trucks  are  moved  to  the  tank-house  where  the  anodes  are 
picked  up,  and  a  group  of  22  forms  the  charge  for  a  tank.  There 
are  420  tanks  in  all,  each  tank  being  in  electrical  connection  with 
the  adjacent  ones.  The  current  passes  through  each  of  the  22 
plates  or  anodes  to  the  cathodes  placed  between,  and  anodes  and 
cathodes  are  2  in.  apart.  Assuming  that  the  anodes  are  2  by  3  ft. 
in  size,  we  have,  in  each  tank,  a  total  area  of  264  sq.  ft.  through 


OF    THE    COMMON    METALS.  435 

which  the  current  is  passing  with  a  density  of  15  amp.  per  sq.  ft. 
We  thus  have  a  total  of  3960  amperes.  The  plates  are  immersed  in 
a  solution  (electrolyte)  of  copper  sulphate  (blue-stone)  in  dilute 
sulphuric  acid,  the  CuSO4  being  15%  (4%  Cu)  and  the  H2S04  10% 
of  the  solution.  The  pressure  in  passing  through  this  2  in.  of 
electrolyte  to  the  cathode  is  0.25  volt.  The  tanks  in  series  would 
therefore  give  a  presure  of  105  volts. 

For  the  cathodes,  so-called  'stripping  plates'  are  prepared  as 
follows:  Certain  tanks  are  reserved  for  this  duty  and  are  provided 
with  insoluble  anodes  of  lead  plates.  Opposite  the  anodes  are 
inserted  greased  copper  plates  upon  which  a  layer  of  copper 
precipitates.  When  the  layer  becomes  1/32  in.  thick  the  plates  are 
taken  out  and  the  sheets  of  copper  are  stripped  from  the  surface,  the 
grease  upon  which  prevents  adhesion.  Some  of  the  sheets  are  cut 
into  strips,  and  made  into  loops,  and  are  riveted  to  the  plates  that 
are  suspended  in  the  regular  tanks.  There  are  23  plates  for  each 
tank  and  these  form  the  beginning  of  the  cathodes. 

One  ampere  deposits  one  ounce  avoirdupois  in  24  hours.  Each 
cathode,  2  by  3  ft.,  increases  in  weight  11  Ib.,  and  all  the  plates  of  a 
tank  248  Ib.  in  24  hours.  The  anodes  correspondingly  decrease  in 
weight  and  become  thinner  by  the  same  amount.  At  the  end  of  two 
weeks  the  cathode,  now  weighing  160  Ib.,  is  removed,  and  another 
starting-sheet  is  put  in  the  place  to  receive  all  the  remaining  copper 
that  can  be  taken  from  the  anode.  Not  all  the  copper  of  the  anode 
is  dissolved.  The  part  of  the  anode  above  the  solution,  and  at  least 
the  skeleton  of  the  plate,  must  remain  to  transmit  the  current. 
Shreds  and  pieces  of  copper  drop  to  the  bottom  of  the  tank,  and 
these  with  the  fragmentary  anode,  still  retaining  15%  of  the 
original  weight,  are  removed,  re-melted,  and  cast  into  new  anodes. 

Current  density. — The  greater  the  density  (in  amperes)  of  the 
electric  current  through  the  anodes,  the  more  rapid  is  the  deposition, 
and  hence  the  smaller  is  the  stock  of  metal  that  needs  to  be  carried 
for  a  given  output.  At  Great  Falls,  Montana,  the  current  density 
is  40  amp.  per  sq.  ft.,  while  at  Perth  Amboy,  N.  J.,  the  density  is 
15  to  17  amp.  At  high  density  there  is  a  liability  of  short-circuiting, 
and  the  cathode-plates  take  on  more  impurity  such  as  arsenic  and 
antimony.  It  commonly  takes  a  month  to  dissolve  anodes.  Hence 
in  a  large  refinery,  a  half-million  to  a  million  dollars  is  tied  up  in 
copper  and  precious  metals  undergoing  treatment. 

Treatment  of  the  slime  or  anode-mud. — Blister-copper,  made 
into  anodes,  is  commonly  97  to  98%  copper.  The  remaining  2  to  3% 
is  largely  substance  insoluble  in  the  electrolyte.  The  insoluble 


436  THE    METALLURGY 

substance  falls  to  the  bottom  of  the  tank,  forming  a  thin  mud, 
valuable  for  the  gold  and  silver  contained.  To  remove  the  mud  the 
tank  is  by-passed.  Being  cut  out  of  the  system,  the  clear  electrolyte 
is  decanted,  and  the  anode  mud  washed  into  a  tight  tram-car  placed 
below  the  outlet-plug  of  the  tank.  The  mud  contains  fragments  of 
copper  of  all  sizes  that  have  dropped  from  the  anodes  when  nearly 
dissolved.  They  are  removed  by  passing  the  mud  through  a  40- 
mesh  screen  into  a  settling  tank.  From  here  it  is  drawn  to  a 
pressure-tank,  then  to  a  filter-press,  and  is  dried  by  steam  to  2% 
moisture  and  broken  up.  The  broken  material  is  then  ready  for 
treatment  in  an  English  cupelling-furnace.  It  is  charged  into  the 
furnace  on  a  molten  lead-bath  into  which  it  melts,  the  lead  taking  up 
the  silver  and  gold.  The  impurity  in  the  dried  material,  such  as 
copper,  arsenic,  antimony,  and  tellurium,  are  partly  volatized,  partly 
absorbed  by  the  litharge-slag  that  forms  under  the  action  of  the 
air-blast  used  in  this  type  of  furnace.  The  lead  having  been  finally 
cupelled,  the  dore  silver  is  left  and  is  cast  into  bars. 

The  scrap  copper,  screened  from  the  anode  mud  as  above 
described,  is  re-melted  and  re-cast  with  other  anode-scrap  into 
anodes,  all  amounting  to  15  to  20%  of  the  copper  treated. 

Another  method  of  recovering  the  precious  metals  consists  in 
digesting  the  anode  mud  in  dilute  sulphuric  acid  which  dissolves 
and  removes  the  copper. 

Purifying  the  electrolyte. — The  best  proportion  of  acid  and 
copper  for  the  electrolyte  is  10%  H2SO4  and  15%  CuS04  (equivalent 
to  4%  Cu).  When  the  copper  exceeds  this  quantity,  the  resistance 
.increases;  hence  copper  is  removed  from  the  circulating  electrolyte 
or  solution  if  in  excess,  to  bring  the  amount  to  the  required  pro- 
portion. The  quantity  of  iron,  arsenic,  antimony,  and  tellurium 
gradually  increases,  and  a  time  comes  when  the  electrolyte  becomes 
foul  with  them  and  the  excess  must  be  removed.  Antimony  can  be 
kept  low  by  the  daily  addition  of  a  small  amount  of  salt,  which 
precipitates  as  an  oxychloride. 

To  purify  the  electrolyte  the  following  method  is  used.  A 
portion  of  the  electrolyte  is  diverted  in  a  constant  flow  to  tanks 
reserved  for  the  purpose  of  purification.  These  have  insoluble  sheet- 
lead  anodes  and  copper  cathodes.  A  strong  current  is  used  so  that 
not  only  is  copper  deposited  but  also  the  impurity.  The  deposit 
collects  loosely  upon  the  copper  plates  and  falls  to  the  bottom  of  the 
tanks.  Every  two  months  the  accumulated  mud,  containing  40  to 
60%  copper,  is  cleaned  out  and  reduced  in  a  reverberatory  refining 


OF    THE    COMMON    METALS.  437 

furnace  to  form  impure  bars  of  copper.  The  purified  electrolyte 
is  returned  to  the  main  system. 

Circulation  of  the  electrolyte. — To  avoid  short-circuiting,  and 
to  increase  the  activity  and  regularity  of  deposition,  the  electrolyte 
is  made  to  flow  or  circulate  through  the  tanks,  entering  the  top 
of  each  tank  near  the  end,  passing  downward  between  the  plates, 
and  finally  rising  and  flowing  away  through  an  overflow  pipe  at 
the  other  end.  After  the  solution  has  flowed  through  two  tanks  in 
this  way,  it  enters  a  launder  that  returns  it  to  the  collecting  or 
sump-tank.  Thus  every  pair  of  tanks  has  an  independent  circulation. 
The  sump  tank  receives  all  the  electrolyte.  Here  the  electrolyte 
is  heated  by  means  of  a  steam-coil  to  40°C.,  the  effect  of  the  warming 
being  to  decrease  the  electrolyte  resistance.  It  is  then  pumped  up 
to  a  distributing  or  stock-tank,  and  thence  once  more  enters  the 
circulation. 

Testing  the  current. — Besides  the  voltmeter  and  ammeter  to  be 
found  at  the  switch-board  in  the  power-house,  it  is  customary  to 
use  a  voltmeter  for  constantly  testing  the  drop  in  potential  between 
the  anodes  and  cathodes.  For  this  a  forked  rod  is  used,  which 
touches  the  two  plates  and  takes  a  small  current  through  a  portable 
voltmeter.  A  slight  drop  of  pressure  indicates  short-circuiting. 

160.     COST  OF  ELECTROLYTIC  REFINING. 

In  the  early  days  of  refining,  costs  were  high  ($20  per  ton),  and 
refineries  charged  $40  per  ton  of  blister-copper  treated.  At  present, 
the  cost  of  refining  98%  copper  anodes  has  been  reduced  to  $4  to  $5 
per  ton,  and  contracts  have  been  closed  at  $7.50  per  ton.  To  this 
cost  is  added  the  charge  for  re-melting,  if  the  copper  comes  to  the 
refinery  in  the  form  of  ingots  that  need  it.  Anodes,  wherever  made, 
should  be  re-melted,  since  casting  them  directly  from  the  converter 
leaves  a  rough  and  porous  product  that  is  undesirable  for  the  elec- 
trolytic tank.  The  cost  of  re-melting  is  generally  stated  at  $5  per 
ton.  The  economy  now  possible  is  the  result  of  cheaper  power  and 
the  use  of  casting  and  handling  machinery,  as  well  as  improved 
operation  in  other  respects. 

Capital  in  plant. — Not  only  is  capital  invested  in  the  buildings 
and  the  equipment  of  the  plant,  but  it  is  required  for : 

(1)  The  stock  of  anodes  in  process  of  treatment. 

(2)  The  stock  of  anodes  awaiting  treatment. 

(3)  The  copper  constantly  contained  in  electrolyte. 

(4)  The  copper  needed  for  the  heavy  conductors  transmitting 
the  current. 


438  THE    METALLURGY 

The  result  of  this  large  demand  upon  capital  is  to  restrict  the 
operation  of  plants  to  places  near  financial  centers,  like  New  York, 
where  cheap  money  is  available,  the  copper  near  the  market,  and 
the  labor  abundant.  These  considerations  outweigh  the  advantages 
of  having  the  plant  near  cheap  water-power. 

161.     REFINING  IMPURE  SPELTER. 

In  general  the  spelter,  as  made  at  the  zinc  furnace,  is  sufficiently 
pure  and  is  used  without  refining.  If,  however,  the  ore  contains 
lead,  the  spelter  will  also  contain  it,  and  this  must  be  removed  by 
refining.  A  small  quantity  of  iron  also  remains  in  the  spelter,  and 
this  must  be  removed.  It  is  observed  that  the  spelter  first  made 
and  expelled  by  distillation  is  purer  than  the  part  obtained  later, 
and  often  the  first  of  the  zinc  is  kept  separate,  molded  into  smaller 
plates,  and  is  put  on  the  market  as  a  special  brand. 

The  principle  of  the  method  consists  in  re-melting  the  spelter 
in  a  reverberatory  furnace  with  a  reducing  flame,  and  letting  the 
molten  bath  stand  until  the  metal  separates  into  layers  according 
to  the  specific  gravity  of  the  different  metals,  the  lower  part  of 
the  bath  consisting  of  a  leady  zinc  and  the  upper  part  of  spelter 
nearly  free  from  lead.  The  lower  layer  is  then  tapped,  or  removed 
otherwise.  The  separation  or  refining  must  be  done  at  a  temperature 
near  the  melting  point  of  zinc  since  the  higher  the  temperature 
the  more  persistently  does  the  zinc  retain  lead.  Under  the  most 
favorable  circumstances  the  lead-content  of  the  spelter  is  reduced 
to  1  to  1.25  per  cent. 

To  refine  spelter,  a  furnace  resembling  that  shown  in  Fig.  182 
and  183  is  used,  but  the  fire  box  is  in  two  parts  and  provision  is 
made  to  charge  the  spelter  close  to  the  bridge.  It  holds  30  tons 
of  spelter  when  full,  and  in  it  10  tons  can  be  refined  in  24  hours. 
The  metals  separate  into  layers.  At  the  bottom  is  the  lead;  the 
iron  forms  with  the  zinc  and  part  of  the  lead  a  difficultly  fusible 
alloy  that  floats  on  the  lead,  and  uppermost  is  the  stratum  of  pure 
zinc.  By  means  of  an  iron  rod  inserted  into  the  bath,  layers  are 
distinguished,  the  zinc  being  soft,  the  iron-lead-zinc  alloy  (called 
'hard  zinc')  being  mushy,  and  the  molten  lead  at  the  bottom  soft. 
The  underlying  lead  is  removed  weekly.  A  cylinder  or  pipe  closed 
at  the  lower  end  is  sunk  below  the  lead  layer.  The  plug  is  then 
knocked  out,  and  the  lead,  rising  in  the  cylinder,  is  ladled  into 
molds.  The  zinc  of  the  top  layer  is  ladled  out  daily  into  molds, 
and  it  retains  1  to  1.25%  lead.  The  hard  zinc  layer  is  removed  when 
opportunity  offers.  To  do  this  the  zinc  is  ladled  out  first,  the  lead 


OF    THE    COMMON    METALS.  439 

is  next  removed,  and  finally  the  mushy  mass  of  ferruginous  metal 
is  removed  with  ladles  perforated  so  that  the  lead  drains  off.  This 
hard  zinc  is  sold  for  the  manufacture  of  Delta  or  Sterro-metal. 

Spelter  produced  in  the  United  States  is  generally  pure  and 
needs  no  refining.  American  high-grade  spelter  contains  only  0.01 
to  0.02%  lead,  and  0.01  to  0.02%  iron.  American  Western  spelter 
ranges  from  0.4%  lead  and  0.02%  iron  in  the  better  brands  to  1% 
lead  and  0.05%  iron  in  the  poorer. 

162.     PARTING  GOLD-SILVER  BARS. 

The  bars  from  reduction  works  contain  gold  and  silver  commonly 
alloyed  with  copper,  but  sometimes  also  zinc  and  lead.  It  is  custom- 
ary to  re-melt  the  bars  and  assay  them,  buying  them  on  the  result 
of  the  assay. 

The  bars  are  parted  in  nitric  or  sulphuric  acid.  Sulphuric  acid, 
being  cheaper,  is  the  acid  commonly  employed. 

Bars  containing  a  large  proportion  of  gold  are  inquartated  by 
melting  them  with  silver  in  order  to  decrease  the  proportion  of 
gold  to  silver  ratio  of  at  least  2.5  to  1,  otherwise  the  acid  fails  to 
attack  the  silver.  In  parting  with  sulphuric  acid  the  copper  should 
be  less  than  10%,  but  in  nitric-acid  parting  more  than  10%  is 
allowed.  To  adjust  this  percentage,  bars  low  in  copper  are  melted 
with  those  high  in  that  metal. 

Nitric-acid  parting. — This  method,  still  practised  at  the  United 
States  Mint,  Philadelphia,  is  an  efficient  way  of  parting,  especially 
on  a  small  scale.  The  bars  having  been  melted  and  proportioned 
as  above  described,  the  molten  metal  is  granulated  by  pouring  it 
into  a  tank  of  water.  The  granulated  metal  is  transferred  to 
porcelain,  glass,  or  platinum  vessels,  and  treated  with  nitric  acid, 
1.20  sp.  gr.,  until  action  ceases.  The  solution  is  allowed  to  settle, 
then  is  decanted,  and  fresh  acid  is  added  for  the  purpose  of  dissolving 
the  remaining  traces  of  silver.  This  is  again  decanted,  and  the  gold 
residue  is  washed  thoroughly  with  hot  water.  It  is  then  ladled  out, 
drained,  dried,  and  mixed  with  a  little  flux,  and  melted  in  a  graphite 
crucible. 

From  the  decanted  silver  solution  the  metal  is  precipitated  by 
adding  common  salt.  The  silver  chloride  thus  formed  is  washed, 
thoroughly  granulated,  zinc  is  added  to  reduce  the  silver  to  metal, 
and  the  zinc  chloride  resulting  from  the  reaction  is  washed  from 
the  precipitated  silver.  The  silver  is  pressed  into  cakes,  melted, 
and  cast  into  ingots  for  bar-silver. 

Sulphuric-acid  parting.— Since  for  this  method  of  parting  there 


440  THE    METALLURGY 

should  be  less  than  10%  copper  present  in  the  gold-silver  alloy,  and 
not  less  than  2g*  parts  silver  to  one  of  gold,  the  bars  to  be 
parted  that  exceed  the  required  limit,  are  so  selected  and  melted 
with  others  as  to  afford  the  required  proportion.  The  melted  metal 
is  cast  into  flat  ingots  and  parted  in  this  form. 

The  ingots  are  placed  in  a  cast-iron  kettle,  covered  with  a  sheet- 
iron  hood  that  is  connected  with  a  chimney  so  that  the  acid  fume 
from  the  kettle  is  carried  away.  Here  they  are  treated  with 
sulphuric  acid  of  full  strength  (66°Be).  When  action  has  ceased 
the  solution  is  allowed  to  settle,  after  which  the  clear  supernatant 
part  is  decanted,  being  drawn  off  by  a  lead-pipe  siphon  into  a  lead- 
lined  precipitating  tank.  The  residue  in  the  kettle  is  treated  six 
or  seven  times  with  fresh  boiling  acid.  In  this  way  the  silver 
completely  dissolves,  the  acid  solution  being  removed  after  each 
treatment.  The  brown  gold  residue  is  finally  boiled  with  water, 
being  heated  and  agitated  by  live  steam  from  a  pipe  inserted  in, 
the  water.  In  this  way  the  gold  is  'sweetened.'  The  residue  is 
removed  from  the  kettle,  dried,  melted  in  crucible  with  a  little  borax 
for  flux,  and  cast  into  a  bar  of  gold  of  999  fine. 

The  acid  solution  from  the  kettles,  which  flows  to  the  precipi- 
tating tank,  is  diluted  with  water,  and  the  silver  is  precipitated  by 
hanging  copper  plates  about  one  inch  in  thickness  in  the  solution. 
The  copper  replaces  the  silver  in  the  acid  solution,  which  becomes 
blue  in  color.  When  precipitation  is  complete  the  clear  solution 
is  decanted,  and  the  cement  silver  at  the  bottom  of  the  tank  is 
washed  with  hot  water  to  remove  the  acid  copper-solution.  The 
'cement'  silver,  or  precipitated  silver,  is  removed  to  a  box,  then 
pressed  into  cakes  or  cheeses  in  a  hydraulic  press.  Thus  compressed, 
it  is  ready  for  melting  in  plumbago  crucibles  after  adding  a  little 
borax  flux.  In  large  establishments  the  silver  is  melted  in  a  small 
reverberatory  furnace  where  it  can  be  conveniently  fluxed,  skimmed, 
and  ladled  into  bars  for  the  market.  The  bars  weigh  35  Ib.  or  500  oz. 
each.  On  refining,  each  bar  is  marked  with  a  number  and  the 
exact  weight,  and  fineness,  and  the  name  of  the  refinery  that  pro- 
duced it. 

163.     REFINING  CAST-IRON  TO  MAKE  WROUGHT-IRON  AND 

STEEL. 

Because  of  the  large  amount  of  carbon  in  cast-iron,  it  is  too 
weak  and  brittle  for  many  engineering  purposes.  Three-foiirths  of 
the  pig-iron  in  the  United  States  is  made  into  steel  or  wrought-iron, 
about  3%  being  manufactured  into  wrought-iron.  Wrought-iron  is 


OF    THE    COMMON    METALS.  441 

used  in  preference  to  steel  for  certain  purposes  because  of  the  welding 
quality,  ductility,  and  toughness  compared  with  bessemer  or  open- 
hearth  steel,  but  for  most  engineering  purposes  steel,  that  now  is 
as  cheap  as  wrought-iron,  is  superior. 

164.     PUDDLING  PIG-IRON  TO  MAKE  WROUGHT-IRON. 

In  this  process  the  pig-iron  is  melted  in  a  reverberatory  furnace 
lined  with  iron  ore,  using  an  oxidizing  flame.  During  the  melting 
there  is  an  elimination  of  silicon  and  manganese  which  are  oxidized 
in  part  by  the  flame  and  in  part  by  the  lining  that  with  the  silicon 
produces  a  slag.  After  melting,  the  heat  is  reduced  and  reaction 
starts  between  the  iron  oxide  of  the  slag  and  the  silicon,  carbon, 
phosphorus,  and  sulphur,  of  the  bath,  whereby  the  impurities  become 
oxidized  and  absorbed  in  the  slag. 

The  removal  of  the  impurities  or  metalloids  leaves  the  metal 
in  the  state  of  wrought-iron,  but  it  is  so  nearly  infusible  that  the 
heat  of  the  furnace  fails  to  keep  the  charge  molten,  and  the  metal 
'comes  to  nature'  or  becomes  pasty.  The  puddler  collects  the  iron 
in  the  furnace  into  several  balls  weighing  125  to  180  Ib.  each,  that 
are  removed,  dripping  with  slag,  and  carried  to  the  jaws  of  a 
squeezer  by  means  of  which  the  slag  is  squeezed  out  of  them.  The 
squeezed  balls  are  sent  to  the  rolls  and  are  rolled  into  bars.  The 
puddled  bar,  called  the  'muck-bar'  is  cut  into  lengths,  and  the 
pieces  are  made  into  bundles,  half  the  bars  being  piled  cross-wise. 
They  are  wired  together  and  heated  in  a  re-heating  furnace  to  a 
welding  heat,  then  rolled  into  bars  of  a  smaller  size  than  the  first. 
The  re-rolled  material  is  known  as  'merchant  bar,'  and  the  effect 
of  the  further  treatment  is  to  eject  more  slag  and  cause  a  cross- 
fibre  structure  in  result  of  the  position  of  the  cross-piled  bars.  A 
sample  of  hand-puddled  bar  has  been  found  to  contain  carbon 
0.296%,  silicon  0.12%,  sulphur  0.134%,  and  phosphorus  0.139  per 
cent. 

165.     STEEL-MAKING   BY   THE   ACID   BESSEMER   PROCESS. 

The  pig-iron  used  in  the  bessemer  process  preferably  contains 
1%  silicon  and  0.5%  manganese,  but  to  make  a  salable  steel,  the 
phosphorus  should  be  below  0.10%  and  the  sulphur  below  0.08%, 
since  neither  element  is  removed  in  the  converter.  If  the  silicon 
is  above  1%  the  large  quantity  of  slag  produced  carries  away  iron. 
If  far  below  1%,  the  charge  does  not  blow  hot.  When  manganese 
is  high  (1.5%),  it  makes  the  charge  sloppy,  the  slag  then  being 
highly  fluid  and  easily  ejected  during  the  blow. 


442 


THE    METALLURGY 


The  converters  in  a  large  plant  are  supplied  from  several  blast- 
furnaces, and  to  insure  a  good  average  pig-metal,  it  is  customary  to 
collect  the  product  of  the  several  furnaces  in  a  single  tilting 
reverberatory  furnace,  called  after  the  inventor  the  Jones  mixer, 
capable  of  holding  300  tons  of  pig-metal.  From  the  mixer  it  is 
drawn  to  the  converters  as  needed,  and  a  regular  supply  is  thus 
insured. 

166.     THE  ACID  BESSEMER  PROCESS. 

The  conversion  is  done  in  an  upright  converter,  lined  as  shown 
in  Fig.  184,  with  a  lining  of  silicious  material  held  together  with 


Fig.    184.      SECTION  OF   BESSEMER  CONVERTER  IN  UPRIGHT   POSITION. 

fire-clay.  Fig.  184  represents  a  converter  in  vertical  section.  It  is 
9  ft.  diam.  by  15  ft.  6  in.  high,  and  is  capable  of  treating  a  charge 
of  15  tons  of  pig-metal.  It  is  swung  on  trunnions,  through  one  of 
which  the  compressed  air  needed  in  operation  enters  to  the  tuyeres 
at  the  bottom.  The  tuyeres  are  19  in  number  and  each  is  provided 
with  12  holes  %  in.  diam.,  or  228  holes  in  all.  The  slag  made  in  a 
converter  is  high  in  silica,  and  has  but  little  effect  on  the  lining, 
so  that  the  lining  lasts  several  months.  The  mouth  of  the  tuyere, 
however,  comes  in  contact  with  the  iron  oxide  formed  during  the 


OF    THE    COMMON    METALS.  443 

blow,  and  hence  this  part  of  the  converter  lasts  only  20  to  25  hours. 
This  bottom  accordingly  is  made  so  that  it  can  be  replaced  by 
another  causing  a  delay  of  20  minutes  in  changing. 

Operation  of  the  converter. — The  hot  converter,  from  which  the 
metal  of  a  blow  has  just  been  poured,  is  placed  in  a  horizontal 
position  and  15  tons  pig-iron  is  poured  into  it  by  means  of  a  ladle 
that  is  brought  from  the  mixer.  When  the  converter  is  in  this 
position  no  metal  can  flow  into  the  tuyeres  and  obstruct  them. 
After  the  metal  is  poured  in,  the  blast  (or  'wind')  is  applied  at  the 
rate  of  25,000  cu.  ft.  per  min.,  the  converter  being  at  this  time 
turned  to  the  vertical  position.  The  blast  now  blows  in  fine  streams 
upward  through  18  in.  of  molten  metal.  Active  oxidation  of  the 
manganese  and  silicon  results  and  in  about  four  minutes  they  are 
oxidized  by  the  oxygen  of  the  air  and  have  become  slag.  The  carbon 
now  begins  to  oxidize  to  CO,  and  this  also  streams  upward  through 
the  metal  and  issues  with  the  air  from  the  mouth  of  the  converter 
in  a  body  of  flame.  After  another  six  minutes  the  flame  shortens 
or  drops,  and  the  operator,  knowing  that  the  carbon  has  been  elimin- 
ated, turns  the  converter  into  horizontal  position,  the  wind  being  at 
the  same  time  shut  off.  In  anticipation  of  this,  a  weighed  quantity  of 
spiegel  iron  or  'spiegel'  has  been  tapped  from  the  spiegel-cupola, 
where  it  is  kept  melted,  into  a  ladle.  The  ladle  is  transferred  by 
the  traveling  crane  and  poured  into  the  converter.  So  great  has 
been  the  heat  evolved  by  the  oxidization  of  the  impurities  of  the 
pig  during  the  ten  minutes  of  the  blow  that  the  temperature  is 
higher  than  at  the  start,  and  we  have  a  white-hot  liquid  consisting 
of  comparatively  pure  metal.  Oxidation-products  remain  in  the  bath, 
and  the  carbon  and  manganese  of  the  charge  tend  to  reduce  these, 
the  unused  carbon  being  in  sufficient  quantity  to  impart  the  desired 
strength  to  the  steel.  Silicon,  which  also  is  introduced,  tends  to 
dispose  of  gas  contained  in  the  metal.  After  the  spiegel  or 
'recarburizer'  has  been  added  and  the  reactions  have  ended,  the 
steel  is  poured  from  the  converter  into  a  ladle.  The  ladle,  after  a 
short  interval,  is  carried  to  a  position  over  the  ingot  molds  into 
which  the  steel  is  to  be  teemed  or  poured.  The  teeming  ladle  is 
'bottom-poured',  that  is,  a  tap-hole  and  plug  are  arranged  in  the 
bottom,  so  that  when  the  ladle  is  brought  over  the  ingot  mold  a 
stream  of  metal  drops  straight  downward  into  it  until  it  is  filled; 
and  so  on,  the  molds  are  filled  successively  until  the  ladle  has  been 
emptied.  The  metal  remains  until  solid,  after  which  the  molds  are 
stripped,  leaving  the  ingots  standing.  The  ingot  is  picked  up, 
and  conveyed  to  a  re-heating  furnace,  and  finally  sent  to  the  rolls  to 


444 


THE    METALLURGY 


be  formed  into  the  shapes  desired  for  market  use.  Fig.  185  illustrates 
graphically  by  curves  the  progress  of  the  reactions,  and  the  elimina- 
tion of  impurities  during  the  blow.  From  it  we  see  the  rate  at 
which  the  easily  oxidized  manganese  and  silicon  are  burned  and  also 
the  carbon,  which  is  but  little  acted  upon  until  these  disappear,  but 
which  after  they  are  gone  oxidizes  rapidly.  The  pig  contains  at  the 
beginning  3.5%  C,  1.0%  Si,  and  0.5%  Mn,  all  being  removed.  The 
recarburizer  adds  to  it,  as  Fig.  185  indicates  1%  Mn,  0.7%  C,  and 
0.15%  Si.  The  manganese  is  added  to  take  from  the  metal  the 
oxygen  absorbed  during  the  blow;  the  carbon  is  to  give  the  steel 


to   II     12    13 

Fig.    185.      ACID    BESSEMER    BLOW,   AMERICAN   PRACTICE. 

the  required  strength  and  hardness,  and  the  silicon  to  dispose  of 
the  gas  contained  in  the  bath. 

167.     THE  BASIC  OPEN-HEARTH  PROCESS. 
During  the   past  fifteen   years  the   bessemer  process   has   been 
gradually  giving  way  to  the  basic  open-hearth  process,  due  to  the 


OF    THE    COMMON    METALS. 


445 


fact  that  low  phosphorus  ore  is  being  exhausted.  Phosphorus 
imparts  the  quality  of  brittleness  to  steel  if  present  in  excess  of 
0.1%,  and,  since  no  phosphorus  is  eliminated  in  the  bessemer  process 
or  in  smelting,  the  iron  ore  supplied  must  be  low  in  phosphorus. 
Suitable  ore  is  called  'bessemer  ore,'  and  ore  having  phosphorus 
above  the  desired  limit  is  called  non-bessemer  ore.  The  limit  may 
be  considered  0.085  Ib.  phosphorus  per  1000  Ib.  of  iron  in  the  ore. 
It  is  claimed  that  for  most  purposes  open-hearth  steel  is  better 
than  bessemer,  but  the  latter  gives  the  most  satisfactory  product  for 
tin-plates,  and  is  well  suited  to  the  manufacture  of  rails.  An  im- 
portant advantage  in  the  basic  open-hearth  process  is  that  it  can  be 
used  for  making  steel  from  pig-iron  and  ore  high  in  phosphorus. 
To  make  steel  in  this  way,  lime  is  added  to  produce  a  basic  slag, 


Fig.   186.     LONGITUDINAL  SECTION  AND  ELEVATION  OF  OPEN-HEARTH 

FURNACE. 


the  hearth  is  lined  with  basic  material  to  withstand  the  action  of 
the  slag,  and  impure  iron  and  scrap  that  contain  phosphorus  are 
used.  To  a  limited  extent,  sulphur  is  removed  by  the  operation. 

Fig.  186  is  a  sectional  elevation,  and  Fig.  187  a  plan  of  a  basic 
open-hearth  furnace,  having  a  hearth  of  30  ft.  6  in.  by  14  ft.  wide. 
With  the  furnace  two  pairs  of  regenerating  chambers  are  connected, 
one  pair  taking  the  heat  from  the  product  of  combustion  or  waste 
gas  after  it  leaves  the  furnace,  and  the  other  pair  pre-heating  the 
producer-gas  and  air  entering  the  furnace.  Every  twenty  minutes 
the  current  is  reversed,  the  escaping  gas  going  to  the  other  pair 
of  regenerators,  while  the  air  and  gas  go  through  the  regenerators 
that  have  just  been  heated  by  the  escaping  product  of  combustion. 
In  this  way,  not  only  is  heat  utilized  for  pre-heating,  but  the  gas 
and  air,  thus  pre-heated,  naturally  give  a  high  temperature  in  the 
furnace.  The  heat  in  fact  is  so  increased  by  successive  reversals, 


446 


THE    METALLURGY 


that  it  could  be  made  to  melt  the  roof  of  the  furnace  itself,  and  indeed 
care  must  be  taken  that  it  does  not  do  this.  As  shown  in  the  plan, 
a  set  of  valves  is  provided.  Certain  valves  are  shown  to  be  open, 
and  others  are  closed,  to  adjust  the  currents  as  desired.  Air  and 
gas  enter  at  the  chambers  at  the  left  and  the  waste  gas  escapes 


Main  Gas  Flue 


i 

« 

1 

Casting  Pit 

*o 

Fig.    187.      SECTIONAL,  PLAN  OF  OPEN-HEARTH   FURNACE. 

through  the  other  pair  to  the  stack.  The  temperature  of  the 
checker  work  near  the  furnace  becomes  1000°C.,  and  near  the  stack, 
400°C.  Thus  the  air  and  gas  entering  the  furnace  become  heated 
to  1000°C.  The  highest  temperature  of  the  furnace  is  1600  to  170()°C. 
The  hearth  of  the  furnace  is  lined  to  the  depth  of  24  in.,  and 


OF    THE    COMMON    METALS. 


447 


carries  a  bath  of  molten  metal  of  that  depth,  its  surface  being 
even  with  the  level  of  the  side  door.  The  lining  is  of  calcined 
dolomite  (CaOMgO)  held  together  with  10%  its  weight  of  anhydrous 
tar.  The  tar  burns  to  a  strong  coke,  that  firmly  unites  the  mass 
into  a  hard  substance.  Pure  magnesite  (MgO)  is  more  expensive 
than  dolomite,  but  it  lasts  longer  and  sometimes  is  used.  It  is  the 
basic  lining  that  gives  the  basic  open-hearth  furnace  the  name.  The 
roof  and  side,  above  the  slag  level,  are  made  of  silica  brick  on 
account  of  the  infusible  nature  of  the  material,  and  it  is  customary 
to  put  in  a  layer  of  neutral  material,  chromite  brick  (FeOCr2O3), 


o  £  Q  iO  hour* 

Fig.    188.      CHEMICAL  CHANGES   IN   BASIC  OPEN-HEARTH   PRESSES. 


between  the  acid  brick  above  and  the  basic  lining  below.  A  basic 
furnace  lasts  350  hours  (18  to  24  weeks)  without  radical  repairs. 

Operation. — In  present  practice  the  charge  for  a  basic  furnace 
consists  of  steel  scrap  (steel  trimmed  in  the  process  of  manufacture, 
old  steel  rails,  and  steel  collected  by  junk  dealers)  ;  of  pig-iron 
containing  less  than  1%  Si,  more  than  1%  Mn,  and  up  to  2%  in  P; 
of  calcined  limestone  (quicklime)  8  to  30%  of  the  charge;  and  of 
iron  ore. 

As  shown  in  Fig.  188,  it  takes  4  hours  to  melt  a  charge,  and  6 
additional  hours  to  complete  the  manipulation,  so  that  in  10  hours 
the  charge  is  ready  to  draw.  During  the  3  to  4-hour  melting  period, 
the  carbon,  manganese,  and  silicon  we  can  see  are  reduced.  The 


448  THE    METALLURGY 

reactions  are  controlled  by  the  melter,  who  sees  that  the  carbon 
is  eliminated  last,  and  if  it  is  oxidizing  too  fast  he  must  'pig  up'  the 
charge  by  the  addition  of  pig-iron  to  increase  the  carbon.  On  the 
other  hand,  if  phosphorus  is  oxidizing  too  fast,  the  oxidation  of 
the  carbon  can  be  hastened  by  'oreing  down'  (adding  iron  ore)  to 
produce  the  following  reaction : 

Fe203  -f-  30  =  2Fe  +  3CO 

If  carbon  is  eliminated  too  soon,  much  iron  becomes  oxidized. 
With  the  oxidation  of  silicon  and  phosphorus  to  silica  and  phosphoric 
acid,  acids  form  with  lime  and  iron  oxide  a  basic  slag  containing 
10  to  20%  -Si02,  5  to  15%  P,  45  to  55%  CaO,  and  10  to  25%  Fe. 
The  slag  does  not  attack  the  basic-lined  hearth,  and  retains  the 
phosphorus  and  the  sulphur,  but  the  CaO  must  be  as  high  as 
possible  for  this,  and  yet  not  so  high  as  to  render  the  slag  infusible. 
After  melting,  active  oxidation  begins,  and  the  bath  boils  by  the 
escape  of  gas.  Upon  the  completion  of  the  operation  the  charge  is 
ready  for  tapping  into  a  50-ton  ladle,  the  metal  filling  the  ladle 
and  the  light  slag  overflowing  and  being  thus  removed.  If  the 
slag  remained,  phosphorus  would  be  reduced  from  it,  upon  addition 
of  the  recarburizer,  and  would  again  enter  the  steel. 

The  recarburizer  is  now  added  to  the  metal  in  the  ladle.  It 
consists  of  charcoal  or  coke  contained  in  a  dozen  paper  bags,  each 
holding  half  a  bushel.  These  are  tossed  into  the  ladle,  then  ferro- 
manganese  is  added,  and  then  a  pound  of  aluminum.  Half  the  fuel 
is  burned,  half  is  absorbed  by  the  steel;  and  the  ferro-manganese 
supplies  the  necessary  manganese  and  silicon  and  a  part  of  the 
carbon.  The  aluminum  absorbs  oxidation  products  from  the  metal. 

The  ladle  is  now  picked  up  by  the  75-ton  traveling  crane,  brought 
to  the  ingot  molds,  and  turned  into  them.  The  further  treatment 
of  the  steel  has  been  described  under  the  'Bessemer  Process.' 

The  furnace  is  patched  or  repaired  where  needed  with  a  mixture 
of  dolomite  or  magnesite  and  tar,  and  is  ready  for  a  new  charge. 
As  shown  in  the  diagram  Fig.  188.  the  resultant  steel  contains 
0.35%  manganese,  0.12%  carbon,  and  still  retains  0.050%  phosphorus 
and  0.020%  sulphur. 

168.  THE  BETTS  PROCESS  FOR  THE  ELECTROLYTIC  RE- 
.  FINING  OF  LEAD. 

The  principle  of  this  process  depends  upon  the  solubility  of  lead 
in  an  acid  solution  of  lead  fluosilicate,  which  is  used  as  an  elec- 
trolyte. The  solution  is  formed  by  diluting  hydrofluoric  acid  con- 


OF    THE    COMMON    METALS.  449 

taining  35%  HF  with  an  equal  volume  of  water  and  saturating  with 
powdered  quartz  according  to  the  reaction : 

2  +  6HF  =  H2SiF6  +  2H20  Cxfe, 


In  the  hydrofluosilicicTead  is  dissolved  until  the  solution  -is- 8% 
lead,  after  which  there  remains  11%  H2SiF6  in  excess. 

The  anodes  are  plates  of  the  base-bullion  to  be  refined,  cast 
2  in.  thick,  resembling  ordinary  copper  anodes. 

The  cathode-sheets  that  receive  the  deposited  lead  are  'stripping 
plates, '  obtained  as  in  the  case  with  copper  cathodes.  They  are  made 
by  depositing  lead  upon  steel  cathode-plates,  prepared  for  use  by 
cleaning,  coating  with  copper,  lightly  lead-plating  them  in  the  tanks, 
and  greasing  with  paraffin.  On  them  is  deposited  the  lead,  and 
when  the  coating  is  of  the  desired  thickness  the  steel  cathodes  are 
removed  from  the  bath,  and  the  lead  coating  or  sheets  are  stripped 
off  for  use  as  cathode.  Another  method  consists  in  casting  the 
cathodes  in  the  form  of  thin  sheets. 

The  anodes  and  cathodes  are  placed  ll/2  to  2  in.  apart  in  the 
tank.  As  in  copper  refining,  the  anodes  are  in  multiple,  and  the 
tanks  in  series. 

The  fall  of  potential  between  anode  and  cathode  is  but  0.2  volt, 
and  the  current  strength  is  15  amp.  per  sq.  ft.  One  ampere  deposits 
31/4  oz.  lead  per  24  hours  following  the  ratio  of  the  atomic  weights 
of  copper  and  lead  which  is  63  to  207. 

In  the  process  the  impurities  remain  as  an  adherent  coating 
on  the  anode,  and  consist  of  the  copper,  bismuth,  arsenic,  gold,  and 
silver.  The  zinc,  iron,  cobalt,  and  nickel  dissolve  in  the  electrolyte. 

As  compared  with  ordinary  refined  lead,  electrolyticalty-refined 
lead  is  pure,  being  practically  free  from  bismuth,  even  when  much 
is  present  in  the  base-bullion,  and  it  is  remembered  that  bismuth 
is  harmful  to  '  corroding  lead. ' 

The  residue  or  anode-slime,  averaging  8000  oz.  or  more  of  silver 
and  gold  per  ton,  is  treated  by  boiling  it  with  sulphuric  acid,  using 
a  steam  pipe  inserted  in  the  solution  to  boil  and  agitate  it  with 
free  access  of  air.  The  washed  residue  is  melted  in  a  small  basic- 
lined  reverberatory-furnace,  the  copper  is  removed  by  using  nitre 
as  a  flux,  and  the  antimony  by  the  addition  of  soda.  The  dore  bars 
finally  obtained  are  parted  in  the  usual  way  with  sulphuric  acid. 


PART  X.     PLANT  AND  EQUIPMENT 


PART.  X.  PLANT  AND  EQUIPMENT. 

169.  PRIMARY  PLANT  AND  EQUIPMENT. 

A  mine  has  a  practical  value  only  when  it  has  been  developed 
sufficiently  to  show  what  the  future  has  to  offer  in  erecting  a 
reduction  works.  Where  the  value  of  the  ore  justifies,  the  beginning 
of  metallurgical  operations  is  kept  simple,  relying  upon  the  skill 
of  the  workman,  and  not  using  a  complicated  plant.  After  knowledge 
is  obtained  in  this  way,  the  introduction  of  an  efficient  plant  may 
be  undertaken.  This  can  be  carried  out  to  the  best  advantage, 
at  times,  by  the  erection  of  a  single  unit,  adding  other  units  when 
the  first  has  been  brought  to  a  practical  success. 

170.     LOCATION  OF  WORKS. 

The  location  of  a  reduction  works  depends  upon  whether  the 
ore  of  a  single  mine  is  to  be  treated  or  a  custom-works  is  to  be 
built.  If  a  mining  company  builds  a  works  to  treat  the  ore  of 
its  own  mine  it  is  usual,  in  order  to  save  the  expense  of  trans- 
portation, to  place  the  plant  as  near  the  mine  as  the  securing  of  a 
suitable  site  and  water-supply  permits.  In  the  case  of  a  smelting 
works,  where  fuel  and  flux  must  be  freighted,  and  where  the  product 
is  to  be  shipped,  then  in  addition  to  the  above  requirements,  a  site 
must  be  considered  with  reference  to  a  railroad. 

A  custom-works,  which  buys  ores  from  different  mines  and 
localities  wherever  it  can,  requires  a  place  as  convenient  as  possible 
to  the  chief  source  of  supply  and  to  the  coke,  flux,  etc.,  that  it 
must  use.  For  such  a  plant,  a  point  should  be  chosen  where  several 
railroads  give  rise  to  competition  in  freight  rates.  Such  a  center 
supports  a  large  population,  and  this  affords  an  abundant  supply 
of  labor. 

The  location  of  a  works  is  also  affected  by  the  cost  of  freight 
on  the  ore  and  on  the  supplies  such  as  fuel.  It  is  also  affected 
by  the  railroad  and  labor  conditions,  the  local  market,  and  the 
capital  or  credit  for  obtaining  money  at  a  low  rate  for  operating. 

Thus  iron  and  steel  manufacture  has  centered  about  Pittsburg, 
Pennsylvania,  because  coke,  coal,  and  natural  gas  is  abundant,  and 
because  a  good  market  is  found  there  for  the  products.  On  the  other 
hand  the  iron  smelter  at  Pittsburg  must  pay  for  freight  from  mine 


454  THE    METALLURGY 

to  furnace,  $2.25  per  ton,  and  must  carry  a  large  supply  of  ore  to 
last  through  the  winter  months  when  navigation  is  closed.  The 
United  States  Steel  Corporation,  the  largest  manufacturer  of  iron 
and  steel  in  the  world,  is  to  erect  a  plant  near  the  iron  ranges 
at  Duluth,  Minnesota,  for  reasons  shown  below. 

Vessels  carrying  iron  ore  to  Lake  Erie  ports  can  return  with 
cargoes  of  coke  or  coal  to  supply  the  Duluth  furnaces.  They  then 
have  a  local  market,  and  it  is  not  necessary  to  'stock  up'  with 
a  winters  supply  of  fuel.  Figuring  roughly  that  2Vi>  tons  coal, 
made  into  coke  and  into  producer  gas,  is  required  to  make  a  ton 
of  steel,  there  is  a  slight  advantage,  as  to  fuel,  in  making  coke  in 
by-product  ovens  at  Duluth,  Minnesota,  and  using  gas  engines 
which  utilize  the  blast-furnace  gases  to  the  best  advantage.  Nearly 
two  tons  of  iron  ore  must  be  sent  to  Eastern  furnaces  to  produce 
this  one  ton  of  steel. 

It  is  seen  from  the  cost  of  producing  zinc,  that  3.5  tons  of 
coal  are  needed  per  ton  of  ore.  Thus  it  is  cheaper  to  convey  ore 
to  fuel,  than  coal  to  the  mine  where  ore  is  produced.  Near  Joplin, 
Missouri,  there  is  ore  and  also  fuel,  we  expect,  therefore,  to  find 
the  zinc  smelting  works  working  there  to  the  best  advantage.  The 
region  is  made  more  favorable  by  the  fact  that  natural  gas  is  to 
be  had  there. 

With  respect  to  silver-lead  works  using  lead  as  a  collector 
of  other  metals,  the  favored  places  have  been  found  to  be  railroad 
centers,  such  as  Denver,  Pueblo,  and  Salt  Lake.  From  12  to  15% 
coke  is  used  in  the  charge  in  such  smelting,  so  that  nearness  to 
coal-fields  is  not  the  all-important  condition.  On  the  other  hand, 
ores  are  available  there  in  proportion  favorable  to  combining  profit- 
ably with  one  another.  The  lead  of  one  ore  and  the  iron  of  another 
being  combined  serve  the  requirements  of  smelting. 

In  treating  ore  by  milling  and  cyaniding,  the  amount  of  fuel 
and  other  supplies  required  is  small,  and  hence  the  natural  place 
for  the  work  is  near  the  mine  that  produces  the  ore,  provided  the 
extraction,  or  recovery  of  the  precious  metals,  is  high.  When, 
however,  the  ore  is  refractory  and  the  recovery  is  low,  it  pays  to 
ship  the  ore  to  smelting  works  that  guarantee  a  high*  extraction. 
The  site. — The  advantage  in  a  side-hill  (or  terraced)  site  as 
related  to  the  level  or  flat  site  has  been  much  discussed.  It  has 
been  claimed  as  an  advantage,  that  the  former  permits  the  ore  to 
advance  by  gravity  from  one  operation  to  the  next,  and  that,  as 
it  becomes  reduced,  there  is  no  necessity  to  again  elevate  it. 

On  the  other  hand  the  flat  site  has  the  following  advantages: 


OF    THE    COMMON    METALS.  455 

(1)  The  first   cost  of  the  works  is  small,   since   grading   and 
retaining  walls,  that  would  be  needed  on  a  side-hill  site,  are  reduced 
to  a  minimum. 

(2)  The  arrangement  can  be  more  convenient,   since  there  is 
no  need,  as  in  a  side-hill  plant,  to  place  the  different  parts  of  the 
plant  in   definite   order  to   obtain  the   required  fall.     When   it   is 
desired  to  expand  the  works,  it  is  possible  to  extend  in  any  direction. 

(3)  Every  square  foot  of  ground  can  be  made  the  equivalent 
to  a  lower  or  an  upper  terrace  or  can  be  left  level.     Hence  the 
parts  of  the  plant  that  must  be  far  apart,  on  a  terrace  site,   can 
be  side  by  side  on  a  level  one.     Ventilation  is  good  and  the  plant 
is  accessible  for  supervision. 

Of  course  on  a  level  site  one  must  use  elevators  or  other  hoisting 
appliances,  but  it  is  seldom  that  we  find  a  terrace  site  where  elevators 
are  not  used.  Certain  products  often  must  be  returned  for  re- 
treatment,  and  the  cost  of  elevating  is  less  than  0.5c.  per  ton. 

Iron  and  steel  plants,  tL^  largest  in  the  world,  are  constructed 
on  level  ground.  The  ore  is  unloaded  direct  from  vessels  to  the 
stock  pile,  using  grab-buckets  holding  5  to  10  tons  each.  If  trans- 
ported in  cars,  the  cars  are  loaded  in  a  similar  manner.  The  custom 
is  to  use  hopper-bottom  cars,  from  which  the  ore  drops  into  the  charg- 
ing bins,  and  thence  by  charge-cars  is  conveyed  to  the  furnace-skip. 
By  the  skip  it  is  hoisted  100  ft.  to  the  furnace-top.  Many  recent 
silver-lead  smelting  plants  occupy  level  sites,  but  the  dumping- 
ground  is  at  a  lower  level. 

For  iron  works  little  attention  is  paid  to  the  location  of  the 
dump.  There  is  no  hesitation  in  sending  the  slag,  if  necessary,  a  mile 
away  by  locomotive  to  be  dumped. 

Mill-sites. — On  the  unclaimed  mineral  lands  of  the  Western 
United  States,  title  is  secured  from  the  general  government  for  a 
mill-site  for  reduction  works,  five  acres  in  extent,  either  in  connection 
with  a  mining  claim  (on  a  theory  that  each  mining  lode  is  entitled 
to  a  mill-site)  or  as  a  site  for  an  independent  or  custom  reduction 
plant.  A  reduction  company,  operating  a  mill,  must  dispose  of  the 
tailing  it  produces,  and  of  the  water  discharged,  not  encroaching 
upon  the  property  of  other  people,  and  it  is  responsible  for  all 
damages.  A  company  must  not  let  tailing  flow  into  a  stream  that,  at 
a  reasonable  cost,  can  be  impounded,  nor  run  into  waters  where 
liable  to  interfere  with  navigation.  The  right  or  custom  of  dumping 
on  the  valueless  land  of  lower  mining  claims  is  general,  except 
that  the  practice  must  do  no  damage  to  the  property  of  owners 
below. 


456  THE    METALLURGY 

A  reduction  company  can  take  up  lands  for  a  ditch  or  flume 
from  unappropriated  public  land,  and  the  claim  can  not  be  interfered 
with  by  later  locators;  but  the  owner  of  such  a  ditch  or  flume  is 
responsible  for  damage  arising  from  breaks  or  overflows. 

The  same  rule  holds  with  respect  to  roads  and  trails.  In  Colorado, 
mining  claims  are  subject  to  the  right-of-way  to  parties  hauling 
ore  over  them,  but  in  other  States  the  location  gives  exclusive 
control,  except  that  a  water,  electric,  or  railroad  company  can 
take  it  under  the  law  of  eminent  domain  by  giving  a  fair  compen- 
sation for  it. 

The  smoke  from  the  smelting  works,  especially  those  treating 
sulphide  ore  in  quantity,  delivers  into  the  atmosphere  many  tons 
of  sulphur-fume  daily,  as  well  as  fine  flue-dust  carried  out  of  the 
stack  by  the  draft.  This  diffuses  through  the  atmosphere  and  is 
carried  by  the  wind  to  trees  and  the  crops  of  the  land.  If  not 
diluted,  it  blights  vegetation,  and  naturally  the  farmers  organize 
to  secure  damage,  or  to  close  the  works.  The  question  of  what  to 
do  to  avoid  the  difficulties  is  a  serious  one,  and  today  when  pyrite 
smelting  and  extensive  roasting  of  sulphide  ores  is  in  practice, 
the  trouble  can  not  be  altogether  overcome.  Thus  far  the  solution 
has  consisted  in  locating  the  works  in  places  where  there  is  little 
vegetation  to  be  damaged,  or  in  discharging  the  fume  into  the 
atmosphere  from  high  stacks.  It  may  be  said  that  the  latter  expedient 
lessens  but  does  not  altogether  obviate  the  difficulties.  The  metal- 
lurgist must  give  serious  consideration  to  the  matter  therefore, 
otherwise,  after  erecting  and  starting  the  operation  of  a  plant, 
he  may  find  that  he  is  compelled  to  close  it,  to  the  ruin  of  the  entire 
enterprise. 

171.     INSTALLATION  OF  PLANT  AND  EQUIPMENT. 

Preliminary  to  building  a  plant  and  operating  a  works,  an 
investigation  is  made  of  the  process,  the  requirements  of  the  plant, 
and  all  limiting  conditions.  It  includes,  besides  the  general  matters 
outlined  above,  the  questions  of  supplies,  markets,  railroad  facilities, 
freight  rates,  sufficient  and  suitable  labor  not  liable  to  strikes,  and 
reliable  civil  conditions  unaffected  by  revolutions  or  oppression 
by  the  government  under  which  the  plant  must  operate. 

Next  comes  the  organization  of  the  operating  company  and 
financing  of  the  enterprise,  or  obtaining  capital  to  build  and  operate 
the  plant  until  it  pays  the  operating  costs. 

Often  the  promoters,  besides  owning  the  mine  for  which  the 
reduction  works  are  built,  have  acquired  the  necessary  real  estate  and 


OF    THE    COMMON    METALS.  457 

the  rights  that  go  with  it.  Provisions  should  be  made  for  access  by 
railroads,  for  the  necessary  trackage,  and  for  the  common  roads  to 
the  plant.  Not  only  must  water  and  power  be  provided,  but  right- 
of-way  for  securing  them.  If  fluxes  are  needed  then  the  proper 
quarries  or  deposits  must  be  found. 

Construction. — Before  beginning  construction,  plans  should  be 
fully  worked  out  by  competent  engineers.  Cost  estimates  are  made 
in  detail,  good  materials  are  accumulated,  and  the  labor-force  is 
properly  organized.  In  the  design  of  the  plant  provision  for 
duplicate  parts  is  made,  so  that  in  case  of  break-down  no  interruption 
of  operation  occurs. 

On  beginning  construction  the  hydraulic  works,  where  needed, 
are  put  in  under  skilled  supervision.  This  includes  the  building 
of  dams,  reservoirs,  the  water-power  plant,  and  transmission  line. 
For  a  long-distance  power-transmission  line  there  may  have  to  be 
sub-stations  and  a  distributing  system. 

Money  must  be  provided  for  the  salaries  of  officers  of  the  company 
that  are  to  receive  pay  during  the  period  of  construction,  and  all 
money  expended  must  be  accounted  for,  and  cost-records  kept  by 
a  skilled  accountant.  The  money  needed  for  legal  expenses,  general 
expense,  traveling  expense,  and  all  expenses  incurred  during  con- 
struction must  be  included. 

Equipment.— This  includes  the  machinery,  furnace-tools,  and 
appliances  used  in  operating,  but  excludes  land,  buildings,  and  track- 
age. Labor-saving  machinery,  when  reliable,  effects  a  saving  in  costs, 
but  it  is  remembered  that  this  saving  must  not  sacrifice  the  efficiency 
of  operation.  The  question  'how  much'  often  arises,  and  we  may 
even  come  to  the  conclusion  that  it  is  not  desirable  (considering 
the  cost  of  installation)  to  put  in  the  labor  saving  appliance. 

172.     HANDLING  MATERIALS. 

The  application  of  machinery  to  the  handling  of  materials  has 
of  late  years  received  much  attention.  It  has  been  rapidly  developed 
because  of  the  resulting  economy  in  labor. 

For  handling  on  one  level,  100  tons  or  less  of  material  daily, 
especially  where  the  ore  is  to  be  distributed  to  various  places,  one 
or  two-wheeled  buggies,  or  barrows,  on  a  good  floor,  have  been 
found  to  be  economical,  elastic,  and  low  in  first  cost.  For  small 
quantities  the  metallurgist  is  not  led  into  installing  machinery,  for 
he  finds  in  practice  that  it  effects  no  saving.  For  large  quantities 
barrows  or  buggies  may  be  used,  or  hand-propelled  tram-cars,  as 
in  mining.  For  still  larger  quantities,  power-propelled  cars  are 


458 


THE    METALLURGY 


used,  that  can  be  handled  also  on  up-grades  and  sent  from  level 
to  level.  The  idea  is  well  carried  out  at  the  Washoe  plant  of  the 
Anaconda  Copper  Mining  Co.,  where  industrial  locomotives  move 
thousands  of  tons  of  material  daily  from  level  to  level.  Indeed  the 


Fig.    189.      VERTICAL  BELT   ELEVATOR. 


OF   THE    COMMON    METALS.  459 

plant  is  an  admirable  example  of  how  a  side-hill  site  becomes  effec- 
tive, where  locomotives  can  be  used. 

Appliances  for  mechanical  handling  are  divided  into  continuous 
machines,  and  appliances  for  intermittent  handling,  such  as  light 
railways  and  cranes. 

Continuous  machines. — These  carry  a  distributed  load,  so  that 
the  sub-structure  upon  which  they  rest  is  light  compared  with  one 
upon  which  the  load  is  concentrated  as  in  a  car.  They  deliver 
material  continuously,  and  no  time  is  lost  in  loading  and  unloading. 
Intermittent  conveying,  on  the  contrary,  if  we  increase  the  load 
of  the  skip  or  bucket,  becomes  slow  and  awkward,  whereas  in  the 
continuous  conveyor  it  is  possible  to  increase  the  capacity  by  widen- 
ing the  conveyor  and  providing  the  correspondingly  increased  feed. 

We  divide  continuous  machines  into  elevators,  conveyors,  and 
conveyor-elevators. 

Elevators  are  used  for  vertical  or  nearly  vertical  lifting.  The 
belt  elevator,  Fig.  189,  is  of  this  type,  and  consists  of  an  endless 


Fig.    190.      STEEL   ELEVATOR  BUCKET. 

belt  having  sheet-steel  buckets,  Fig.  190,  attached  by  flat-headed 
elevator-bolts  at  18-in.  intervals.  To  allow  for  the  stretching  of 
the  belt,  the  upper  pulley  shaft  is  carried  in  take-up  boxes,  by 
which  the  shaft  can  be  raised.  It  is  generally  preferable,  however, 
to  use  the  take-up  boxes  for  the  lower  pulley-shaft.  The  lower 
pulley  is  enclosed  in  a  boot,  the  ore  delivering  into  the  buckets  at 
the  rising  side,  shown  at  the  left.  Ore,  not  caught  by  the  buckets, 
falls  into  the  boot  and  is  there  scooped  out  by  buckets.  The  boot 
has  a  hinged  drop-bottom,  so  that  it  can  be  cleaned  out  when  desired. 
The  drop-bottom  is  of  particular  advantage  for  elevators  used  in 
a  sampling  mill  in  cleaning  up  between  samples.  The  speed  of 
the  belt  is  275  to  300  ft.  per  min.  to  insure  that  the  material  delivers 
to  the  discharge  spout  by  centrifugal  action,  as  the  buckets  pass 
over  the  top  pulley.  At  this  speed  a  belt-elevator,  having  buckets 
8  in.  wide,  delivers  7  tons  per  hour,  and  a  10-in.  bucket-elevator, 
18  tons  per  hour.  The  head  and  boot-pulleys  are  30  in.  diam.  and, 


460 


THE    METALLURGY 


as  well  as  the  belt,  are  made  1  to  2  in.  wider  than  the  buckets.  The 
whole  is  inclosed  in  a  wooden  housing  to  prevent  the  escape  of  dust. 
Fig.  191  represents  a  single-strand  endless-chain  elevator.  The 
chain  is  carried  by  head  and  foot  sprocket-wheels  with  sprockets 
spaced  to  take  links  of  the  chain.  In  this  case  the  'take-ups'  are 


Fig.    191.      SINGLE-STRAND   END- 
LESS-CHAIN  ELEVATOR. 


Fig.    191.      SINGLE-STRAND   ENDLESS- 
CHAIN   ELEVATOR. 


carried  by  the  boot,  As  is  the  case  with  the  belt-elevator,  the 
velocity  should  be  sufficient  to  insure  an  efficient  discharge  from  the 
buckets ;  but  chain  elevators,  because  of  the  numerous  joints,  are 
not  well  suited  to  run  at  a  high  velocity. 

Fig.  192  is  a  double-strand  endless-chain  elevator  having  two 
head-pulleys  arranged  so  that  the  buckets  discharge  into  a  spout 
between  them.  In  this  way  the  elevator  can  be  run  at  the  low 
velocity  suited  to  the  type.  To  take  up  the  slack,  the  shaft  of  the 
second  head-pulley  is  carried  by  horizontally  moving  take-up  boxes. 


OF    THE    COMMON    METALS.  461 

Either  of  the  endles-chain  elevators  can  be  housed,  in  wood  as  in 
Fig.  189,  or  in  a  sheet-iron. 

Conveyors  are  used  for  the  horizontal  transfer  of  materials, 
and  can  be  modified  easily  to  carry  up  an  incline.  Of  all  conveyors, 
the  belt-conveyor  is  most  widely  used.  To  give  it  capacity  it  is 
troughed  by  running  on  pulleys  that  raise  the  edges  of  the  belt 
forming  a  shallow  trough  (See  Fig.  193).  The  simplest  form  is  an 
endless  belt  running  over  end-pulleys,  the  load  being  fed  at  one 
end,  delivering  into  a  chute  or  into  a  bin  at  the  other.  The  conveyor 
carries  a  load  not  only  on  a  level,  but  on  as  steep  as  24°  incline. 


Fig.    193.      CONVEYING  BELT  WITH  TRIPPER. 

The  capacity  is  large  and  the  conveyors  are  simple  and  durable. 
A  12-in.  belt,  traveling  at  the  rate  of  150  to  350  ft.  per  min., 
delivers  10  to  35  tons  per  hour.  A  24-in.  belt,  traveling  at  the 
extreme  velocity  of  600  ft.  per  min.,  has  a  capacity  of  250  tons  per 
hour  of  crushed  ore,  and  requires  6  hp.  per  100-ft.  length,  the  power 
needed  varying  with  the  length  of  the  belt.  When  the  ore  ascends 
an  incline  we  add  the  power  for  lifting  the  load.  It  is  desired  at 
times  to  deliver  the  ore  into  bins  situated  at  different  points  along 
the  belt.  This  is  accomplished  by  using  the  movable  tripper  shown 
in  Fig  193,  which  also  shows  the  belt  loaded  with  ore.  To  discharge 
the  ore  the  belt  goes  around  the  upper  pulley,  as  shown,  then  around 
a  second  one  just  below,  and  continues  the  course  to  the  front  end- 
pulley.  The  ore  shoots  from  the  belt  into  spouts,  that  deliver  on 
either  side  of  the  track  upon  which  the  tripper  moves.  Indeed  the 
tripper  is  sometimes  made  to  travel  continuously  back  and  forth 


462  THE    METALLURGY 

from  end  to  end  of  a  long  bin  at  the  rate  of  200  ft.  per  minute,  evenly 
distributing  the  ore  and  bedding  it.  The  bin  when  full  supplies 
the  furnace,  and  the  ore  is  of  an  even  constitution  throughout. 

The  worm  or  screw-conveyor. — This  is  convenient  for  delivering 
crushed  ore  short  distances,  and  it  thoroughly  mixes  the  ore  con- 
veyed. Fig.  194  represents  a  screw-conveyor  delivering  ore  from 
the  trough  along  which  it  has  been  conveyed,  into  another  at  right 
angles.  The  ore  drops  from  the  first  to  the  second,  and  is  conveyed 
by  the  screw  in  the  second,  shown  at  the  left.  The  bottom  of  the 
trough  is  lined  with  smooth  sheet-steel  bent  half  round  to  conform 
to  the  worm  or  screw.  A  screw  conveyor  is  shown  in  Fig.  Ill 
illustrating  a  dry-crushing  silver-mill.  The  disadvantages  of  this 


Fig.    194.      SCREW   CONVEYER    (QUARTER  TURN). 

type  of  conveyor  are,  that  much  power  is  needed,  and  that  the 
ore  grinds  on  the  conveyor  resulting  in  wear.  It  is  well  adapted 
to  carrying  a  moderate  and  continuous  supply  of  ore  short  distances. 

Endless-chain  conveyors, — These  are  much  used,  since  they  con- 
vey ore  not  only  on  a  level,  but  vertically  if  necessary.  Being  entirely 
of  metal  they  successfully  convey  hot  materials. 

Fig.  195  represents  an  endless-chain  push-conveyor,  consisting  of 
a  series  of  plates  or  'flights'  attached  to  a  double  endless-chain 
carried  at  each  end  by  sprocket-wheels  like  the  double  endless-chain 
elevator  Fig.  192.  The  ore,  drawn  from  any  desired  storage-bin  as 
shown  in  the  figure,  is  pushed  up  an  incline  by  the  moving  flights 
in  a  fixed  steel-lined  trough,  and  is  taken  by  a  double-strand  endless- 
chain  elevator  to  a  floor  above.  If  desired,  slides  may  be  provided 
in  the  bottom  of  the  trough.  When  the  slide  is  opened,  the  ore 
drops  into  the  desired  bin  beneath. 

Sometimes  in   place   of  flights,   a  continuous  series   of  buckets 


OF    THE    COMMON    METALS. 


463 


or  trays  is  used.  These  overlap  so  that  the  ore  can  not  drop  between 
them.  They  operate  upon  the  principle  of  the  Howden  pig-casting 
machine,  Fig.  126.  Indeed  the  conveyors  lend  themselves  to  a  great 
variety  of  applications,  as  the  examination  of  a  catalogue  of  elevating 
and  conveying  apparatus  will  show.  The  chief  draw-back  to  them  is 
that  they  have  numerous  joints  to  wear,  and  that  the  troughs,  flights, 
or  buckets  are  subjected  to  serious  wear.  They  must  run  slower 
than  the  belt-conveyor. 

In  Fig.  38  and  39,  the  Edwards  roasting-furnace,  we  have  an 
example  of  a  swinging  push-conveyor.  The  flights  are  bladed  and 
so  hinged  from  the  vibrating  carrying-beam  as  to  swing  over  the  ore 


Fig.    195.      ENDLESS-CHAIN  PUSH   CONVEYOR. 

in  the  conveying  trough  on  the  backward  motion,  but  to  push  the 
ore  along  when  moving  forward.  As  is  seen,  the  bottom  of  the 
trough  is  provided  with  slides  to  deliver  the  ore  where  it  is  needed. 
Vibrating-trough  conveyors. — These  are  of  a  simple  type,  consist- 
ing of  a  sheet-steel  flat-bottomed  trough,  50  ft.  long  by  2  ft.  wide, 
supported  by  spring  legs,  and  receiving  a  throw  of  1  in.  from  an 
eccentric  making  300  rev.  per  min.  The  conveyors  can  be  arranged 
to  deliver  into  one  another,  in  series,  a  distance  of  500  ft.  if  desired. 
Owing  to  the  inclination  of  the  spring  legs,  the  trough  rises  in 
the  forward  motion  and  drops  in  the  backward  motion,  so  that  the 
material  is  propelled  to  the  discharge  end.  The  conveyors  are 
simple,  inexpensive,  and  need  few  repairs.  A  conveyor,  having  a 
trough  24  in.  wide,  has  a  capacity  of  20  to  25  tons  per  hour.  When 
provided  with  a  double  or  false  bottom,  the  upper  one  being  a 
screen,  the  conveyor  can  be  made  an  effective  screening  apparatus. 


464  THE    METALLURGY 

Hoists. — Of  these,  the  commonest  about  reduction  works  is  the 
platform  elevator,  which  takes  buggies,  wheelbarrows,  or  tram-cars 
from  floor  to  floor.  It  may  have  a  platform  of  a  size  (6  by  6  ft.) 
to  receive  two  cars  or  wheelbarrows  at  a  time,  and  it  raises  a  one- 
ton  load  60  ft.  per  min.  They  are  often  run  in  balance,  but  it  is 
better  to  have  two  independent  counterweighted  platforms.  Neces- 
sarily, time  is  lost  in  loading  and  unloading,  so  that  the  estimate 
of  the  capacity  is  25  tons  hourly. 

Fig.  122  shows  an  iron  blast-furnace  having  a  platform-hoist, 
and  Fig.  121,  one  having  a  skip-hoist  adapted  to  mechanical  charging. 
The  capacity  of  the  skip  is  2  to  5  tons  of  ore  to  half  the  quantity 
of  coke,  and  they  are  run  in  balance.  It  takes  34  seconds  actual 
time  for  raising,  dumping,  and  returning  the  skip  to  pit ;  but  the 
total  time  including  the  waits  is  4  minutes,  this  furnishing  the 
supply  to  a  furnace  producing  350  to  500  tons  of  pig-iron  daily  from 
a  total  burden  of  1150  to  1650  tons. 

Industrial  railways  and  tram-tracks. — We  already  have  referred 
to  the  use  of  these  in  industrial  work.  Where  an  industrial 
locomotive  can  be  used  it  is  possible  to  convey  on  up-grades,  taking 
advantage  of  side-hill  grades,  or  of  trestles.  Trestles  are  used 
about  level  sites,  and  by  means  of  them  ore  can  be  conveyed  in 
hopper-bottom  cars  and  dumped  into  storage  or  charge-pockets.  Fig. 
137  shows  an  electric  power  system  with  a  motor  for  handling  a 
31  cu.  ft.  or  3!/2-ton  slag-car,  as  used  at  the  United  Verde  Smelting 
works,  Jerome,  Arizona. 

Grabs. — In  large  establishments  hoisting  rigs  are  used  that  are 
provided  with  large  clam-shell  buckets  or  grabs.  They  take  5  to  10 
tons  of  ore  at  a  time,  and  are  used  for  unloading  vessels,  and  for 
transferring  ore  to  stock-piles  for  storage,  or  to  the  furnace  storage 
bins  as  desired.  It  is  noticed  that  the  movable  frames  or  bridges  are 
made  heavy  to  carry  the  large  loads  safely. 

The  traveling  crane. — Fig.  143  gives  in  elevation  a  traveling  crane 
as  commonly  used.  It  is  for  handling  ladles  and  converters  as 
described  under  copper-converting  and  for  bessemer  and  open- 
hearth  practice.  For  handling  materials  inside  a  building  it  is 
coming  into  general  use.  The  cranes  are  operated  by  electricity, 
and  move  in  any  direction,  horizontally  or  vertically,  over  the  floor 
of  the  building  commanded  by  them,  and  they  avoid  obstacles  on 
the  floor.  Provided  with  large  mushroom-shaped  electro-magnets, 
they  are  now  used  to  unload  pig-iron  or  handle  steel  sheets  weighing 
a  ton  or  more,  and  by  using  them  no  time  is  lost  as  in  older  methods, 
in  passing  chains  around  objects  to  be  lifted. 


PART  XI.     COMMERCIAL 


PART  XI.     COMMERCIAL. 

173.     KINDS  OF  WORKS. 

There  are  two  classes  of  metallurgical  works;  those  constructed 
at  mines,  and  custom-works. 

A  mine-works  gets  the  ore-supply  from  the  mine  of  the  company. 
In  such  case,  it  does  not  pay  for  the  ore  it  receives,  and  the  operation- 
cost,  being  only  that  of  reducing  the  ore,  is  low. 

A  custom-works,  that  includes  the  purchase  of  ores  among  the 
items  of  cost,  has  here  a  serious  item  of  expense.  This  is  especially 
applicable  in  the  case  of  the  treatment  of  precious-metal  ores.  The 
stock  of  ore  that  must  be  carried  depends  upon  the  distance  from 
the  mines,  and  upon  the  certainty  of  the  supply.  In  the  case  of 
the  Lake  Superior  iron-ore  supply,  for  example,  a  stock  must  be 
accumulated  by  the  beginning  of  the  winter  to  supply  the  furnaces 
until  the  opening  of  navigation  in  the  spring,  a  period  of  six  months. 
The  silver-lead  and  copper  smelting-works  of  the  Rocky  Mountain 
States  carry  a  supply  to  last  two  to  six  weeks.  We  have  already 
alluded,  under  'Electrolytic  Refining,'  to  the  capital  locked  up  by 
the  process.  On  the  other  hand,  at  a  mine-works,  the  supply  can 
be  replenished  as  used,  so  that  provision  is  needed  only  for  one  to 
two  days'  running,  and  often  for  but  half-a-day's  run. 

174.     ORGANIZATION    OF    A    METALLURGICAL    COMPANY. 

Metallurgical  operations  on  a  commercial  scale  require,  gener- 
ally, the  organization  of  a  company,  or  if  the  company  is  already 
organized,  the  establishment  of  a  department  to  provide  the  addi- 
tional function. 

Where  a  metallurgical  company  is  to  be  organized,  the 
promoters  or  organizers  obtain  a  charter,  or  articles  of  incor- 
poration, from  the  State  in  which  they  desire  to  incorporate.  They 
next  hold  a  meeting  at  which  they  receive  the  property  that  is  t<» 
be  taken  over  by  the  company,  adopt  a  set  of  by-laws  for  the 
guidance  of  the  company,  and  elect  the  directors  that  are  to  manage 
the  affairs. 

The  directors  proceed  to  the  election  of  the  corporate  officers 
of  the  company  from  their  number.  The  officers,  of  a  small 


468  THE  METALLURGY 

company,  are  the  president,  the  vice-president,  the  secretary,  and 
the  treasurer.  The  directors  may  appoint  from  their  number  a 
managing  director,  or  they  appoint  a  manager  from  the  outside  to 
have  charge  of  the  affairs  of  the  company. 

In  outlining  the  organization  of  a  company  undertaking 
metallurgical  works,  the  manager  should  be  guided  by  the  following 
rules : 

(1)  He  should  see  that  a  supreme  authority  is  provided  over 
all  action  to  be  taken,  and  should  carefully  and  fully  outline  the 
authority  and  responsibility  of  each  position,  making  the  duties 
of  each  conform  to  the  capability  of  the  party  holding  it.  To  do 
this  he  must  avoid  making  any  person  subordinate  to  two  or  more ; 
should  place  the  authority  and  responsibility  together;  should 
distribute  the  work  and  the  duties  not  to  overburden  nor  to  under- 
load; and  should  arrange  the  positions  so  that  promotion  can  come 
from  them.  While  the  manager  gives  his  chief  attention  to  the 
commercial  or  business  affairs  of  the  company,  he  generally  appoints 
a  superintendent  to  attend  to  the  technical  affairs  of  the  plant. 

The  organization,  under  the  charge  of  the  manager,  may  include 
(1).  the  supply,  (2)  the  operating,  (3)  the  accounting,  and  (4) 
the  selling  department. 

(1)  The    supply    department    attends    to    the    purchase,    and 
delivery   of  ore,   fuel,   and  flux,   and  to  the   care   and   issuing   of 
chemical  and  general  supplies. 

(2)  The  operating  department  has  to  do  with  all  that  pertains 
to  the  reduction  or  manufacture  of  the  ore  into  metal  (the  winning 
of  the  metal  from  ore)    or  to  refining  metals  to  bring  them  into 
marketable  form,  and  has  control  of  the  operating  forces,  consisting 
of  the  foremen  (and  men  under  them),  the  repair  force  (consisting 
of  mechanics  and  their  helpers  who  keep  the  plant  in  repair  and 
put   in  the  needed  improvements),   and  the  laboratory   or  assay- 
office  force. 

(3)  The  accounting  department  attends  to  the  accounting,  pay- 
roll, cost-keeping,  and  the  distribution  of  costs. 

(4)  The  selling  department  attends  to  the  disposal  and  sale 
of  the  product  of  the  works.      By-products,  in  process  of  further 
treatment,  are  not  here  included. 

175.     THE  PURCHASE  OF  ORES. 

Smelting  works  purchase  ores  of  every  kind  according  to  the 
requirements,  provided  the  ores  are  sufficiently  valuable  to  pay  for 


OF    THE    COMMON    METALS.  469 

treatment.     They  are  purchased  according  to  the  content  and  by  a 
pre-arranged  schedule. 

176.     IRON  ORES. 

These  are  purchased  by  guarantee  on  the  part  of  the  shipper 
that  they  will  come  up  to  a  given  standard  that  generally  is  based 
upon  the  percentage-content  in  natural  condition,  thus  including 
the  contained  moisture.  For  Lake  Superior  ore  the  prices  for  1908 
at  Lake  Erie  ports,  per  long  ton  (2240  Ib.)  were: 

Old  Range  bessemer,  55%  iron  base $4.50 

Old  Range  non-bessemer,  51.5%  iron  base.  . .  .   3.70 

Mesabi  bessemer,  55%  iron  base 4.25 

Mesabi  non-bessemer,  51.5%  iron  base 3.50 


The  Old  Range  ores  come  from  the  iron  ranges  on  the  south  side 
of  Lake  Superior,  and  command  a  higher  price  because  of  the  better 
mechanical  condition.  The  Mesabi  ores  are  soft,  friable,  and  carry 
much  fine,  which  makes  flue-dust.  When  smelting  such  ore.,  in  ratio 
of  85%  soft  to  15%  hard  ore,  as  much  as  6%  flue-dust  or  dirt  is  made. 
On  a  non-bessemer  ore,  as  it  varies  from  this,  a  premium  of  8.349c. 
is  paid  for  each  per  cent  iron  over  the  guarantee,  and  a  penalty  or 
deduction  is  made  of  8.349c.  down  to  50%  iron,  of  12.523c.  down 
to  49%  iron,  and  a  double  penalty  down  to  48%  iron.  Below  48%, 
the^penalty  becomes  18c.  per  unit. 

On  bessemer  ore,  provision  is  made  for  a  premium  only  in  case 
the  ore  exceeds  the  guaranteed  55%  iron. 

177.     ORES  USED  IN  SILVER-LEAD  SMELTING. 

A  schedule  for  the  purchase  of  silver-lead  ores  in  Clear  Creek 
and  Gilpin  counties,  Colorado,  is  here  given,  dated  Feb.  1,  1905,  f.  o. 
b.  cars  at  Denver,  based  on  the  dry-weight  of  the  ore.  The  ton  used 
is  the  short  ton  of  2000  pounds. 

Dry  tailing  and  concentrate. — Gold,  $19  per  ounce,  if  0.05  oz.  or 
more  per  ton;  silver,  95%  of  the  New  York  quotation  on  the  day 
of  assay,  if  1  oz.  or  more  per  ton. 

Copper,  dry-assay  (wet  less  1.5  units), 

Per  Unit. 

5%  or  less $1.25 

Over  5%  and  including  10% 1.50 

Over   10% 1.75 

10%  silica  basis,  lOc.  up;  5%  zinc  basis  30c.  up. 


470 


THE    METALLURGY 


Treatment 

Gross  value  of  ore.  per  Ton. 

Not  over  $35  per  ton  .....................   $3.50 

Over  $35  and  including  $80  per  ton  ........     4.00 

Over  $80  per  ton  ........................     5.00 

Upon  lots  containing  less  than  7  tons  .......     5.00 

Dry-silicious  and  copper-bearing  ore.  —  Gold,  silver,  and  copper 
are  paid  for  as  in  the  schedule  for  concentrate. 

Treatment  charge  $8  on  a  40%  silica  basis,  5c.  down  and  lOc. 
up  to  a  maximum  charge  of  $11  on  ores  not  exceeding  $25  gross 
value;  5%  zinc  limit  30c.  up. 

Lead  ore.  —  Gold  $19.50  per  oz.,  silver  95%  of  the  New  York 
quotation  on  the  day  of  assay,  copper  $1  per  unit  dry  (wet  less 
1.5  units)  when  ore  assays  2%  or  over  wet;  10%  zinc  basis,  50c.  up. 

NEUTRAL    SCHEDULE. 


Lead  Inclusive 
Per  Cent. 
5  to  10 

Per  Unit 
Cents. 
25 

Treatment 
Charge. 

$800 

10  to  15    

25 

700 

<  15  to  20    

25 

5.00 

20  to  25 

25 

400 

25  to  30 

30 

400 

30  to  35                .    .  . 

30 

300 

35  to  40    

30 

250 

40  to  45    

32 

2.00 

45  to  50 

35 

200 

Over  50  . 

40 

2.00 

FLAT    SCHEDULE. 


Lead  Inclusive 
Per  Cent. 
5  to  10 
10  to  15 
15  to  20 
20  to  25 
25  to  30 
30  to  35 
35  to  40 
40  to  45 
45  to  50 
Over  50  . 


Per  Unit 
Cents. 
25 
25 
25 
25 
30 
30 
30 
32 
35 
40 


Treatment 

Charge. 

$12.00 

10.50 

8.50 

6.50 

6.00 

4.50 

3.00 

2.00 

2.00 

2.00 


OF    THE    COMMON    METALS.  471 

Neutral  basis,  lOc.  up  or  down.  The  schedule  most  favorable  for 
the  shipper  to  be  used. 

Oxidized  irony  ores. — Gold  and  silver  are  valued  as  in  the 
schedule  for  concentrate.  Lead  25c.  per  unit  for  5%  or  over. 
Treatment  charge  $2  on  a  neutral  basis,  lOc.  per  unit  up. 

Lead  concentrate. — Gold  $19  per  oz.  if  0.05  oz.  or  over  per  ton ; 
silver  and  copper  as  in  lead  ores.  Silica  basis  10%,  lOc.  up ;  5%  zinc 
limit,  30c.  up. 


Lead  Inclusive 
Per  Cent. 
5  to  10    

Per  Unit 
Cents. 
25 

Treatment 
Charge. 

$4.75 

10  to  15        

24 

4.00 

15  to  20    

30 

3.50 

20  to  25    

32 

3.25 

25  to  30    , 

35 

3.25 

Upon  concentrate  assaying  over  30%  lead  apply  either  the  neutral 
schedule  or  the  flat  schedule  as  under  'Lead  Ores,'  whichever  favors 
the  shipper. 

Copper  ores  were  formerly  purchased  upon  the  result  of  the  fire- 
assay,  which  experience  shows  to  be  about  1.5%  low.  It  was 
considered  that  the  fire-assay  expressed  what  was  to  be  obtained 
from  the  ore  when  treated  on  a  full  scale  in  the  furnace.  If  analysis 
shows  that  the  percentage  of  SiO2  is  equal  to  that  of  Fe,  the  iron 
present  is  considered  able  to  flux  the  silica,  and  the  ore  is  called 
self-fluxing  and  is  said  to  be  neutral,  the  base  (iron)  neutralizing 
the  acid  (silica).  If  more  silica  than  iron  is  present  then  the 
difference  in  percentage  is  called  the  silica  excess,  and  the  reverse 
when  iron  is  in  excess.  Zinc  is  detrimental  to  smelting,  and  if  an 
ore  has  more  than  5%,  a  charge  is  made  against  the  ore  or  a  penalty, 
as  it  is  called,  is  exacted  by  the  smelting  company.  Flat  prices 
are  paid  for  the  lead,  that  is,  they  remain  the  same  whether  the 
price  of  lead  rises  or  falls.  Concentrate  (the  granular  product 
made  in  dressing  or  concentrating  ore)  containing  no  lead  but 
consisting  of  iron  or  copper  sulphide,  is  called  dry.  The  value  is  in 
the  contained  silver  and  gold.  Lead  concentrate  is  valued  also 
for  the  contained  lead. 

As  an  example  of  the  use  of  the  above  schedule,  let  us  take  an 
ore  containing  SiO2,  14%  ;  Fe,  6%  ;  Zn,  11%  ;  Mn,  4%  :  S,  10%  ;  Pb, 
21%  ;  with  60  oz.  Ag  and  0.2  oz.  Au  per  ton.  The  ore  is  evidently 
a  lead  ore,  and  we  will  figure  it  on  both  the  neutral  and  the  flat 
schedule  given  under  'Lead  Ore.' 


472  THE    METALLURGY 

We  find  a  silica  excess  of  4%  over  the  iron  and  manganese.    The 
excess  of  zinc  over  10%  is  one  per  cent. 

On  the  neutral  schedule  we  have : 

Gold,  0.2  oz.  at  $19.50 $  3.50 

Silver,  60  oz.  at  95%  of  62c.  (New  York  quotation) 35.34 

Lead  21%  (21  units)  at  25c 5.25 


Total  metal  value $44.49 

Deducting  treatment,  $4  -f  (4%  Si02  excess  at  lOc.)  +  (1% 

Zn  excess  at  50c.)   4.90 


Net  returns  to  the  shipper  f.  o.  b.  Denver $39.59 

On  the  flat  schedule  we  have : 

Total  metal  value  as  before $44.49 

Deducting  $6.50  treatment  +  50c.  zinc  penalty 7.00 


Net  returns  to  shipper  f.  o.  b.  Denver $37.49 

Therefore  the  neutral  schedule  is  more  favorable  to  the  shipper  and 
is  the  one  used. 

178.  SCHEDULE  FOR  COPPER  ORES. 

Low-grade  copper  sulphide. — Ores  figured  f.  o.  b.  Tacoma,  Wash- 
ington, 1907,  per  short  ton  dry-weight.  Gold  $20  per  ounce,  silver 
95%  the  New  York  quotation,  day  of  assay,  copper  3c.  per  pound  less 
than  New  York  quotation. 

Treatment 
Charge. 

When  net  value  is  less  than  $5  per  ton $3.25 

"      "   more  than  $5  per  ton 3.45 

While  the  above  schedules  give  an  idea  of  prices  on  which  to 
base  estimates  of  the  value  of  ores,  most  buying  is  done  by  special 
arrangement,  and  time-contracts  are  made  for  the  delivery  of  ore. 
The  skilled  buyer,  well  informed  in  the  details  of  the  business,  is 
apt  to  have  an  advantage  over  the  shipper,  and  can  make  profitable 
contracts  for  his  company.  Often  there  is  but  one  buyer,  who  has 
much  his  own  way,  but  the  seller,  who  can  show  to  the  buyer  that 
he  is  well  posted,  makes  better  terms. 

Other  supplies. — With  the  exception  of  ores,  fuel,  and  fluxes, 
supplies  are  kept  in  a  store-room,  and  issued  by  the  supply  depart- 
ment only  on  a  written  order  from  the  foreman  or  other  responsible 
party  that  needs  them.  In  this  way  is  known  where  the  supplies 
are  to  be  distributed  on  the  cost  sheets.  An  account  is  kept  of  all 


OF    THE    COMMON    METALS.  473 

supplies  received  and  issued,  so  that  from  it  can  be  learned  how 
much  and  when  to  order  to  keep  properly  stocked.  It  is  detrimental 
to  the  business  to  so  run  out  of  supplies  as  to  cause  delays. 

Much  knowledge  is  required  in  the  purchase  of  supplies,  to  buy 
when  prices  are  low  or  on  a  rising  market,  but  only  in  small  quantity 
on  a  falling  market,  and  to  obtain  the  best  discounts,  but  care 
is  taken  to  avoid  the  purchase  of  inferior  goods. 

179.     THE  OPERATING  DEPARTMENT. 

The  choice  of  men  to  operate  a  plant  is  often  the  important 
condition  to  the  success  of  the  enterprise. 

The  superintendent  not  only  must  be  informed  as  to  the  actual 
technical  operations,  but  he  must  know  how  to  arrange  and  organize 
his  forces.  He  should  be  able  to  handle  men  effectively,  and  must 
possess  tact,  discretion,  and  firmness.  He  must  be  strict  and  just, 
and  able  to  encourage  as  well  as  to  drive.  He  is  often  the  metal- 
lurgist as  well  as  the  superintendent,  and  has  direct  management  of 
the  furnaces  and  the  metallurgical  machinery.  When  thing  go 
wrong  it  is  he  that  is  called  upon  to  correct  them,  at  whatever  hour 
of  the  day  or  night.  If  a  furnace  is  in  bad  condition  or  a  machine 
out  of  order,  he  is  responsible.  If  all  is  going  smoothly  his  duty 
may  be  light,  but  when  troubles  occur,  or  when  the  company  is 
losing  money,  his  work  is  hard.  If  he  fails  in  adjusting  difficulties, 
no  excuse  is  accepted.  He  must  succeed  or  must  resign.  Much 
of  his  success  depends  upon  his  subordinates,  and  first  in  importance 
among  them  we  must  place  the  foremen. 

In  early  times  the  success  of  the  operation  depended  upon  the 
skill  of  the  workmen.  Today,  manual  skill  must  be  supplemented 
by  intellectual  skill  to  insure  the  certain  and  exact  operation  of 
the  plant.  At  first,  the  puddler  worked  up  his  charge  of  iron  by 
hard  labor.  Today  he  operates  levers  and  watches  the  progress  of 
the  successive  operations.  In  place  of  muscular  effort  the  workman 
now  must  supply  intelligent  direction. 

The  payment  for  the  work  of  supervision  and  control,  as  for  the 
office  force,  superintendent,  assayer,  chemist,  and  foremen,  is  made 
monthly.  In  certain  cases  an  additional  premium  is  paid.  In  plants 
where  the  exhaustion  of  a  mine  or  of  a  mining  district  may  cause  the 
closing  of  a  works,  premiums  can  hardly  be  assured.  The  premium 
method,  however,  has  been  found  to  give  good  results,  especially 
where  rated  on  tonnage,  or  in  the  case  of  the  superintendent,  on  the 
lessening  of  costs.  There  is  a  decided  difference  between  the  in- 
centive of  a  man  who  draws  his  salary  in  any  case,  and  one  who 


474  THE    METALLURGY 

knows  that  his  compensation  increases  with  his  diligence  and  effort 
to  practice  economy. 

Engineering  is  the  art  of  doing  with  one  dollar  what  ordinary 
man  may,  after  a  fashion,  do  with  two ;  hence  a  competent  man,  who 
is  a  trained  expert,  is  necessary  to  best  results,  and  where  the 
margin  of  profit  is  small,  he  is  absolutely  essential.  When  the 
price  of  the  metal  being  produced  is  high,  and  the  margin  of  profit 
is  large,  tonnage  is  more  important  than  economy. 

The  capacity  and  efficiency  of  a  plant  depends  on  the  intelligence 
and  reliability  of  the  men  in  charge  of  the  different  departments, 
and  they  are  chosen  with  these  qualities  in  view.  The  low-grade 
labor  is  used  where  hard  routine  work  comes  in,  the  high  where 
judgment  is  essential.  While  low-grade  labor  may  be  faithful,  it 
is  stupid  and  liable  to  blunders ;  when  trained,  however,  if  faithful, 
it  becomes  reliable.  The  employment  of  cheap  men  for  supervising 
costly  machines  is  offset  by  the  loss  of  time  or  by  actual  disaster. 

Qualities  of  a  foreman. — A  foreman  often  is  a  man  that  has  been 
advanced  from  a  subordinate  position  in  the  works.  He  may  be 
selected  from  among  several  applicants  from  the  outside.  In 
quizzing  an  applicant,  note  the  names  of  his  former  employers,  ask 
him  if  he  'gets  on'  well  with  his  men,  if  he  scolds  them,  if  he  has 
been  threatened  by  them,  or  if  he  treats  them  with  consideration. 
A  foreman  must  not  be  thought  to  be  a  'good  fellow'  (an  'easy 
fellow')  by  his  men.  He  may,  himself,  sometimes  lead  off  in  the 
work  and  show  his  men  how  to  do  things,  but  in  general  he  has 
enough  to  do  in  seeing  that  they  are  all  busy. 

The  foreman  not  only  must  watch  the  progress  of  the  work,  but 
plan  ahead,  to  be  sure  that  everything  is  provided  and  ready  at 
hand  as  needed.  In  such  work  the  foreman  need  not  be  always  so 
strict  a  disciplinarian.  On  routine  work,  and  in  order  to  drive, 
especially  when  he  has  many  men  to  handle,  he  must  be  just,  having 
reserve  of  manner,  but  looking  to  the  welfare  of  the  men  while  at 
work. 

The  chemist,  or  assayer. — The  chemist  or  the  assayer  is  called 
upon  for  results  regarding  the  operation  of  the  plant,  and  these 
he  must  furnish  with  promptness  and  accuracy.  Besides,  in  his 
routine  work,  his  skill  and  value  is  found  in  investigation  under 
the  director  and  advice  of  the  metallurgist  or  superintendent. 

The  construction  force. — The  various  mechanics  (blacksmith, 
carpenter,  engineer,  and  machinist)  not  only  have  repairs,  but  also 
new  construction  to  attend  to,  under  the  personal  direction  of  the 
superintendent,  who  may,  where  the  work  requires,  employ  a  drafts- 


OF    THE    COMMON    METALS.  475 

man  or  constructing  engineer  to  attend  to  construction.  It  is  a 
rule,  in  case  of  a  break-down  or  other  emergency,  when  the  general 
foreman  in  charge  of  operating  needs  aid,  that  this  work  has  pre- 
cedence, and  other  work  must  be  dropped  to  expedite  it. 

Skilled  and  unskilled  labor. — The  ten-hour  labor  at  the  reduction 
works  is  largely  unskilled  or  common  labor.  These  laborers  are 
called  'outside  men'  or  'roustabouts.'  They  do  the  work  requiring 
the  use  of  pick  and  shovel,  such  as  unloading  cars,  handling  the 
products  of  the  works,  and  assisting  in  the  construction  or  repairs. 

Skilled  laborers  working  in  shifts  of  eight  to  twelve  hours,  called 
also  'inside  labor'  receive  higher  pay  per  hour  than  common  laborers. 
The  pay  varies  according  to  the  skill  needed.  These  men  are 
responsible  for  the  successful  performance  of  the  duties  given  them, 
and  they  are  expected  to  work  until  they  are  relieved  by  their 
partners  on  the  following  shift,  or  until  the  foreman  provides  some- 
one to  fill  their  places. 

For  keeping  discipline,  and  to  prevent  slackness  in  the  work, 
certain  rules,  the  result  of  long  experience,  have  been  laid  down 
for  the  guidance  of  the  men.  These  are : 

The  men  must  be  promptly  at  work,  and  must  work  the  full  time, 
or  (for  inside  hands)  until  relieved. 

For  ten-hour  men  the  working-day  is  from  7  a.  m.  to  12  m.  and 
from  1  p.  m.  to  6  p.  m.  in  summer,  while  in  winter  the  noon-hour 
is  shortened  to  30  minutes,  and  the  day  ended  at  5  :30  p.  m. 

Inside  men  must  be  on  hand  the  entire  time  of  their  shift,  and 
must  eat  their  luncheon  when  there  is  time  while  watching  their 
work.  Charge  wheelers  must  keep  up  the  supply,  even  when  the 
furnace  is  running  fast,  but  may  rest  at  intervals,  and  are  not  called 
upon  to  do  other  work  than  sweeping  up  at  their  own  places  before 
going  off  shift.  The  inside  man  can  leave  when  relieved  by  his 
partner,  but  must  wait  for  his  partner  if  delayed  in  arriving.  If 
the  latter  fails  to  appear  the  foreman  provides  another  man,  who 
then  holds  the  place,  the  absent  man  losing  it  unless  he  has  a  good 
excuse,  or  if  sick,  he  is  required  to  notify  the  foreman  by  message, 
who  then  provides  a  man  in  his  place.  When  the  first  man  desires 
to  return  to  work  he  must  notify  the  foreman  one  shift  in  advance, 
so  that  the  substitute  is  not  put  out  of  a  shift  for  which  he  has 
come  prepared. 

When  men  are  sick  on  shift,  if  not  too  seriously,  they  should 
be  held  if  possible  until  the  end  of  the  shift.  It  is  impressed  upon 
them  that  it  is  detrimental  to  the  work  for  them  to  leave,  and  that 
it  is  difficult  at  short  notice  to  get  someone  to  fill  their  place. 


476  THE    METALLURGY 

Men  must  obey  orders,  and  flat  disobedience  is  followed  by 
discharge,  irrespective  of  who  the  man  may  be.  Otherwise  discipline 
is  weakened. 

Be  strict  but  just.  It  helps  in  discipline  to  discharge  the  poorest 
man  occasionally,  and  if  much  time  elapses  without  this  you  may  be 
sure  that  you  are  becoming  less  strict. 

Do  not  entrust  men  to  do  work  without  supervision  and 
inspection.  They  do  it  wrong,  or  become  careless  when  they  realize 
that  they  are  not  watched. 

In  certain  respects  metallurgical  work  differs  from  that  of  other 
industries.  The  work  is  carried  on  by  a  crew  who  work  together, 
each  man  having  a  part  of  the  operation  to  depend  on  him  alone. 
All  are  directed  by  the  foreman  who  sees  to  the  regular  operation 
of  the  mill  or  furnace.  Thus  we  have  about  a  silver-lead  blast- 
furnace, the  feeder  and  his  helper,  the  weigher,  and  the  wheelers 
or  trammers  who  bring  in  the  stock,  all  being  men  that  work  on  the 
feed-floor.  At  the  slag  or  lower-floor  there  is  the  furnace-man,  the 
tapper,  and  the  pot-pushers.  The  men  on  both  floors  are  called 
inside  men.  They  work  in  shifts  of  8  to  12  hours,  and  one  crew 
replaces  another.  The  men  are  by  no  means  paid  an  equal  wage. 
The  furnace-man  or  feeder,  for  example,  is  paid  more  than  the  others, 
and  the  pay  varies  with  the  skill  and  knowledge  required. 

Provision  is  often  made  for  the  care  of  the  men  in  case  of  sickness. 
A  charge  of  one  dollar  per  month  holds  against  every  man,  and 
this  entitles  him  to  medical  attendance  and  care  at  a  hospital  in 
case  of  accident.  If  a  man  has  worked  five  days  in  any  given  month, 
the  one  dollar  is  deducted  from  his  pay  for  hospital  dues. 

While  accidents  occur,  especial  care  is  taken  by  the  foreman,  for 
the  safety  of  his  men.  Men  are  apt  to  grow  careless  and  must  be 
warned,  or  if  disobedient,  discharged.  Any  carelessness  on  the  part 
of  the  foreman  renders  the  company  liable  for  damages. 

The  tendency  at  modern  mills  and  metallurgical  works  is  to 
reduce  the  amount  of  labor  per  ton  of  output,  this  being  done  by 
the  use  of  mechanical  appliances,  so  that  the  labor  of  the  individual 
becomes  less  strenuous,  less  trying  to  the  strength  and  endurance, 
and  yet  better  paid.  In  result,  the  workman  becomes  more  able 
and  is  more  desirous  of  retaining  his  job,  and  his  services  are  more 
reliable.  A  large  mill  is  run  with  less  labor,  proportionally,  than 
a  small  one.  An  easily-treated  ore  requires  less  labor,  since  there 
are  few  operations  and  the  skill  and  care  is  less. 

Eight-hour  shifts  need  one-half  more  men  than  twelve-hour  shifts. 

Steam-power  requires  more  labor  than  water-power,  the  latter 


OF   THE    COMMON   METALS.  477 

often  requiring  only  an  occasional  adjustment  of  the  supply-gate. 

In  dealing  with  Mexican  employees  the  American  engineer  must 
be  scrupulously  honest.  Honesty,  there  as  elsewhere,  is  expected 
of  him,  and  no  training  nor  circumstances  alters  the  restraint  he 
expected  to  put  upon  himself.  The  mining  and  metallurgical 
engineer  is  intrusted  with  the  control  of  large  enterprises  and  the 
care  of  precious  metals,  and  his  success  depends  upon  his  responsi- 
bility. 

In  starting  a  new  reduction-works,  a  list  of  the  places  and 
occupations  of  the  men  should  be  prepared,  and  the  men  that  are 
engaged  are  quickly  assigned.  All  men  thus  engaged  should  be 
questioned  as  to  their  qualifications,  and  given  places  that  they  can 
fill.  When  a  new  works  is  to  be  started,  men  skilled  in  the  operation 
and  duties  apply  at  the  plant  and  often  are  willing  to  accept  common 
labor  awaiting  the  starting  of  the  works.  In  this  way  they  may  be 
retained  until  needed. 

Quality  and  cost  of  labor. — American  labor  is  characterized  by 
intelligence,  energy,  and  responsibility,  as  compared  with  Mexican 
and  much  foreign  labor.  Mexicans  that  have  been  trained  to  duties 
often  make  excellent  inside-men.  When  well  bossed  or  directed,  they 
do  well,  though  often,  from  being  insufficiently  fed,  their  physical 
endurance  is  low.  They  need  watching  to  prevent  them  from  idling 
their  time,  and  they  should  be  carefully  directed,  since  they  are  apt 
to  work  at  disadvantage  on  account  of  their  lack  of  thought.  Treat 
them  like  grown  children,  have  patience  with  them,  but  be  judicious 
and  strict.  An  employer  of  Mexican  labor  may  find  a  man  that 
excels  in  his  work.  There  is  a  danger  that  he  may  be  spoiled,  by 
raising  his  wages.  In  directing  Mexican  labor,  the  familiar  style, 
not  the  formal  or  polite  one  of  the  books,  is  to  be  used.  Mexican, 
and  in  a  less  degree,  other  labor  is  disposed  to  pilfer.  To  such  an 
extent  is  this  true  that  in  Mexico  a  tool  or  other  portable  article 
can  hardly  be  laid  down,  without  danger  of  it  being  stolen. 

Men  must  be  held  responsible  for  the  tools  that  are  served  to 
them,  returning  them  at  night  to  be  locked  up.  It  is  also  well  to 
brand  tools.  In  the  spring,  when  prospectors  everywhere  are  start- 
ing into  the  hills  for  their  summer  occupation,  the  winter's  job  is 
left  at  the  reduction-works,  and  tools  needed  in  prospecting  may 
be  secretly  taken. 

In  determining  the  rate  of  wages  in  a  new  mining  district  in 
Mexico,  it  is  well  to  start  with  low  pay,  as  the  people  have  been 
accustomed  to  it.  It  is  easy  to  raise  wages  later  if  advisable,  but 
difficult  to  reduce  them  when  a  certain  rate  has  been  paid.  Among 


478  THE    METALLURGY 

Mexicans,  whose  tastes  are  simple,  and  wants  easily  supplied,  the 
result  of  paying  high  wages  is  to  give  more  'laying-off'  time,  and 
this  results  in  inconvenience  at  the  works.  Monthly  payments 
interfere  with  the  steady  operation  of  the  works,  because  at  the  time 
of  the  monthly  payment  the  men  are  disposed  to  'lay-off'  to  spend 
their  money.  To  overcome  the  difficulty  two  methods  have  been 
tried.  One  is  that  of  daily  payments,  by  which  a  man,  who  spends 
his  money  when  he  gets  it,  has  only  sufficient  to  supply  his  daily 
wants  and  those  of  his  family,  and  none  remaining  for  drunkenness 
or  gambling.  The  other  system  is  to  pay  a  wage  to  which  is  added 
a  premium  that  increases  with  the  time  worked.  It  is  paid  at  the 
end  of  the  month  if  the  man  works  through  the  month,  but  other- 
wise not.  This  tends  to  keep  him  steadily  at  work. 

180.     DUTIES  AT  A  LARGE  GOLD  STAMP-MILL. 

The  men  needed  will  be :  Foreman,  night  foreman,  pipe-fitter, 
two  engineers,  mill- wright,  two  firemen,  one  head  amalgamator,  four 
amalgamators,  two  oilers,  two  feeders,  two  laborers. 

The  foreman  has  general  supervision  of  the  mill,  and  looks  after 
the  handling,  cleaning,  and  retorting  of  all  amalgam  collected. 
The  amalgamators  dress  the  chuck-blocks  and  plates,  and  keep 
them  in  good  condition.  They  set  tappets,  regulate  the  water  supply, 
and  make  renewals.  The  feeders  attend  to  the  uniform  feeding 
of  the  batteries,  and  assist  the  amalgamators  in  renewals  and  at  the 
clean-up.  A  good  feeder  is  a  valuable  man  about  a  mill.  The  vanner- 
men  attend  to  the  vanners,  or  concentrating  tables.  They  must  be 
men  with  experience,  and  commonly  should  first  serve  at  the  vanner 
as  'sulphide-pullers.'  The  crusher-men  feed  the  crushers  with  the 
mine-ore  as  it  comes  to  the  mill.  Oilers  oil  the  machinery.  Sulphide- 
pullers  remove  the  concentrate  or  sulphide  from  the  vanner  boxes, 
and  store  it  for  shipment.  Engineers  run  the  power-plant,  and 
have  charge  of  the  firemen.  Firemen  fire  the  boilers  and  remove 
the  ashes.  Coal  passers  wheel  in  coal  from  the  coal  pile  to  the 
boilers. 

On  repairs  there  are  carpenters,  with  laborers  to  help  them.  In 
repairs  on  the  vanners  there  is.  a  special  vanner-man  to  assist. 

181.     THE  ACCOUNTING  DEPARTMENT. 

This  department  takes  charge  of  the  accounts  and  transactions, 
makes  up  the  pay-roll,  and  attends  to  the  payment  of  the  men.  It 
has  also  to  collect  and  collate  costs. 

The  costs  are  prepared  for  three  purposes,  as  follows : 


OF    THE    COMMON    METALS.  479 

(a)  To  enable  the  owner  of  the  property  to  judge  as  to  the 
value  and  the  efficiency  of  management. 

(b)  To  add  to  the  economy  of  management  by  showing  where 
economies  can  be  practised  and  leaks  stopped. 

(c)  To  prevent  dishonesty. 

Costs  are  divided  into  flat,  or  prime-costs,  and  general  expense, 
sometimes  called  fixed  charges.  The  flat-cost  varies  directly  almost 
with  the  tonnage,  while  general  expense  remains  almost  constant. 

Cost  may  vary  as  follows  :  Flat-costs  (per  ton  of  ore  put  through) 
diminishes  as  the  output  increases.  The  percentage  of  labor-cost 
to  general  expense  is  at  a  minimum  in  years  of  average  activity. 
It  increases  in  prosperous  times  because  tonnage  is  more  important 
than  close  saving,  per  ton,  and  it  increases  in  dull  years  because 

general  expense  must  be  divided  by  a  smaller  tonnage. 

• 


The  flat-costs  include  labor,  fuel,  operating,  and  supplies. 

General  expense  includes  cost  of  management  and  superin- 
tendence, office  force,  insurance,  indemnities  and  damages,  hospital 
and  medical  expenses,  taxes  and  rates,  and  costs  of  selling  the 
product. 

In  connection  with  the  matter  of  costs  there  are  two  further 
considerations  which  have  a  bearing  on  the  returns  of  the  capital 
invested.  These  are  depreciation  of  plant,  and  interest  on  the 
invested  capital.  A  plant  will  depreciate  at  the  rate  of  10  to  15% 
annually,  so  that  even  if  kept  in  repair,  it  will  have  become  obsolete 
and  worn  out  at  the  end  of  ten  years.  If,  however,  a  company  sets 
aside  from  the  earnings  an  amount  sufficient  to  put  in  the  needed 
improvements,  and  so  keeps  the  plant  up  to  date,  then  depreciation, 
except  for  the  buildings,  may  be  disregarded.  In  case  this  is  not 
done,  then  dividends  must  be  sufficient,  not  only  to  pay  interest 
on  the  capital  invested,  and  a  profit  besides  because  of  the  risk 
involved,  but  also  to  meet  depreciation.  Otherwise  the  investor 
is  not  recovering  his  capital  and  the  compensation  or  interest  due. 

The  following  are  costs  often  not  considered:  Expressage  on 
gold  or  silver  bars  to  the  mint  or  market.  Costs  of  selling,  generally 
a  percentage  on  the  gross  amount  of  the  sale  or  the  metal  costs. 
Interest  on  bonds  and  a  sinking  fund  to  meet  the  payment  of  the 
principal  when  these  mature  —  an  expense  that  must  first  come  out 
of  the  profits  before  the  stock-holders  are  paid  a  dividend.  Home- 
office  expense  due  to  the  fact  that  the  office  may  be  in  a  large  city 
where  the  stock-holders  live,  distant  from  the  works.  The  object 
of  such  an  office  is  that  the  corporate  officers  and  stock-holders 


480  THE    METALLURGY 

may  be  in  touch  with  the  property  in  which  they  have  invested. 
Royalty,  when  a  percentage  must  be  paid  for  the  use  of  a  process. 

182.     LABOR  COSTS. 

In  a  40-stamp  silver  amalgamation  mill  having  24  pans  and  a 
capacity  of  150  tons  of  ore  per  24  hours  the  inside  labor  for  that 
time  was: 

Four  pan-men  (12  hours)  at  $4 $16 

Two  helpers    (12  hours)    at   $3 6 

Ten  tank-men  (12  hours)   at  $3 30 

$52 

The   quoted   prices  prevail   in   California   and   in   other  high-price 
camps  in  the  Rocky  Mountain  region. 

Men  needed  and  the  cost  of  labor  for  a  single  silver-lead  blast- 
furnace plant  for  24  hours,  having  two  12-hour  shifts  are  as  follows : 

On  feed  floor :     Two  feeders  at  $2.50 $  5.00 

Two  feeder's  helpers  at  $1.80 3.60 

Six  charge-wheelers  at  $2.40 10.80 

Two  weighers  at  $2.40 4.80 

On  slag  floor :      Two  furnace-men  at  $2.50 5.00 

Two  tappers  at  $1.80 3.60 

Four  pot-pushers  at  $1.80 7.20 

Two  engine-runners  at  $3.50 7.00 

Two  foremen  at  $4.25 8.50 

Ten-hour  men :     Two  dump-men  at  $1.50 3.00 

Four  sampling  mill-men  at  $1.50 6.00 

One  sampling  foreman 2.50 

Mechanics :  One  blacksmith 3.00 

One   carpenter 3.50 

$73.50 

The  above  is  based  on  a  wage  of  $1.50  per  day  for  common  labor, 
and  for  a  minimum  force,  that  must  be  increased  when  putting  in 
improvements,  or  for  emergencies.  If  they  can  be  profitably  em- 
ployed, it  is  well  to  have  men  to  supply  vacancies  caused  by  sickness, 
accidents,  or  men  leaving. 

The  following  is  the  labor  cost  at  the  Treadwell  stamp-mill,  (240 
stamps)  Douglas  Island,  Alaska. 


OP    THE    COMMON    METALS.  481 

Inside  men : 

One  foreman  at  $150  per  month $  5.00 

Four  amalgamators  (12  hr.)  at  $90  per  month. . . .  12.00 

Eight  feeders  (12  hr.)  at  $70  per  month 18.64 

Four  vanner-men  (12  hr.)  at  $65  per  month 4.34 

Two  sulphide-pullers  (10  hr.)  at  $2.00 4.00 

Two  sulphide-shovellers  (10  hr.)  at  $2.00 4.00 

Two  engineers  (12  hr.)  at  $2.50 5.00 

Two  foremen  (12  hr.)  at  $2.50 5.00 

Two  coal  passers  (10  hr.)  at  $2.00 4.00 

Four  crusher-men  (10  hr.)  at  $2.25 9.00     $79.66 


Repairs : 

One  carpenter  (10  hr.)  at  $4 $  4.00 

One  vanner-man  (12  hr.)  at  $1  per  month 3.34 

One  laborer  (10  hr.)  at  $2 2.00 

$  9.34 


$90.00 

To  these  labor  costs  is  added,  for  board  and  lodging  furnished 
by  the  company,  an  actual  cost  of  at  least  50c.  per  man  daily. 

183.     PROFITS. 

Profits  from  the  operation  of  a  plant,  whether  independent  or 
connected  with  a  mine,  may  be  defined  as  the  difference  between 
the  total  gross  costs  and  the  returns  on  the  metal-product  sales. 
The  profits  may  be  increased  either  by  better  extraction  (recovery) 
from  the  ore,  by  economy  of  treatment  due  to  methods  permitting  a 
saving  in  labor,  supplies,  or  fuel,  and  by  faster  running,  by  which 
the  output  is  increased. 

The  profit  of  a  custom  works,  which  buys  ores  outright,  is  the 
difference  between  the  charge  made  for  treatment  and  the  actual 
cost  of  treatment.  In  the  early  days  of  smelting  at  Leadville, 
Colorado,  charges  of  $60  per  ton  were  made,  the  actual  cost  being 
$20,  and  the  profit  $40.  Today  this  margin  is  as  low  as  $1  to  $2  per 
ton  because  of  competition.  Another  profit  is  made  on  extraction. 
The  works  makes  a  deduction  for  losses  in  extraction.  If  they 
extract  more  than  this,  the  difference  is  a  clear  gain  to  them.  Thus 
they  pay  for  95%  of  the  silver  in  the  ore,  but  if  they  recover  as 
they  generally  do,  98%,  then  the  3%  difference  is  a  profit. 

We  will  take,  as  an  example,  the  profits  that  may  be  expected 
from  the  following  ore : 


482  THE    METALLURGY 

Treatment  charge  per  ton $6.25 

Actual   cost  "     "    4.50 

—        $1.75 

Gain  in  extraction  of  silver $0.25 

"     "  "          "    gold 0.10 

"     "  "•         "  lead    0.35 

0.70 
This  shows  a  total  profit  per  ton,  from  both  resources.  .$2.45 

In  milling,  this  calculation  remains  the  same  whether  the  ore  is 
highly  silicious  or  not,  but  in  smelting  the  silicious  ore  would  make 
more  slag  that  would  contain  more  of  the  metal,  and  this  would 
diminish  the  gain  in  extraction  and  perhaps  convert  it  into  loss. 

The  following  figures  represent  the  profits  of  a  company  owning 
a  mine,  the  Robinson  company,  on  the  Rand,  South  Africa : 

Gold  recovered  at  the  stamps $20.50 

Gold  recovered  by  cyaniding 5.70 


Total  recovery $26.50 

Cost  of  mining. $6.55 

Cost  of  milling 0.98 

Cost  of  cyaniding 0.97 

$8.50 
Net  profits  per  ton .  $17.70 

184.  THE  SELLING  DEPARTMENT. 

Selling  the  product  of  the  works  is  not  difficult.  There  is  a  ready 
market  for  the  commercial  metal,  but  advantage  can  be  taken  of  the 
market,  and  there  are  companies,  aside  from  the  producing  company, 
that  can  be  employed  to  perform  this. 

185.  FUEL  AND  METAL  MARKET. 

In  examining  market  quotations  of  metals  we  must  understand 
clearly  the  kind  of  ton  meant.  The  short  ton  (2000  Ib.)  is  used  in 
the  Western  United  States  for  ore  and  metal.  For  coal,  iron  ore, 
pig  iron,  and  steel  the  long  ton  (2240  Ib.)  is  understood,  particularly 
in  the  Eastern  States  in  the  wholesale  trade.  In  England,  copper, 
tin,  and  spelter  are  weighed  and  sold  by  this  ton.  On  the  continent 
of  Europe,  in  Mexico,  and,  in  fact,  wherever  the  metric  system  is  in 
use,  the  metric  ton  of  2204  Ib.  is  taken,  and  this  approximates  to 
the  long  ton. 

In  the  United  States,  New  York  is  the  chief  market  for  metals, 


OF    THE    COMMON    METALS.  483 

and  the  sale  of  ore  and  of  metal  is  based  on  the  prices  there.  The 
quotations  of  other  markets,  as  San  Francisco,  St.  Louis,  and  London, 
are  also  often  given.  For  coal,  coke,  iron,  and  steel  other  centers 
are  understood.  Referring  to  a  technical  publication,  as  for  example 
the  Mining  and  Scientific  Press,  of  San  Francisco,  or  The  En- 
gineering &  Mining  Journal  of  New  York,  we  find  the  quotations 
published.  Taking  as  an  example  the  quotations  of  September  1908, 
the  necessary  explanations  are  as  follows : 

Anthracite  coal,  New  York  market. — Schedule  prices  are  $4.75 
for  broken  and  $5  for  egg,  stove,  and  chestnut.  Steam-size  prices 
are  unchanged,  pea  $3.25@3.50,  buckwheat  No.  1,  $2.35@2.50,  buck- 
wheat No.  2  or  rice,  $1.60@2,  barley,  $1.35@1.50.  All  prices  are 
per  long  ton  f.  o.  b.  (free  on  board)  New  York  harbor  points. 

Run  of  mine  anthracite  coal  is  sent  to  breakers,  large  buildings 
containing  the  necessary  crushing  machines  and  revolving  screens, 
to  sort  or  screen  the  coal  to  the  sizes  above  specified.  The  smaller 
sizes,  called  'steam  sizes',  are  cheaper,  and  are  utilized  in  burning 
under  steam  boilers. 

Bituminous  coal,  New  York  market.— Good  grades  of  steam-coal 
are  quoted  at  $2.50@2.65,  with  the  poorer  qualities  selling  around 
$2.40.  At  the  mine,  bituminous  coal  is  quoted,  f.  o.  b.  cars,  for  run- 
of-mine,  85c.,  %-in.  gas-coal,  90c.,  and  slack  at  50e.  The  %-in  gas- 
coal  is  obtained,  by  screening  the  slack  through  a  bar-screen  having 
%-in.  spaces.  The  coal  that  drops  through  the 'spaces  is  called  slack. 
The  difference  between  the  prices  at  the  mine  and  at  New  York 
represents  chiefly  the  cost  of  freight  to  New  York. 

Coke,  Pittsburg1  market. — At  the  coke  ovens  the  prices  f.  o.  b. 
are  for  furnace  coke,  $1.65@1.85,  and  for  foundry  coke,  $2.10@2.25. 
To  be  well  suited  for  cupola  use  coke  should  be  firm  and  in  larger 
pieces  than  otherwise  is  required. 

Iron  ore,  market  at  Lake  Erie  ports,  as  Loraine,  Cleveland,  Con- 
neaut,  and  Ashtabula. — Old  Range  bessemer,  $4.50,  non-bessemer, 
$3.75,  Mesabi  bessemer,  $4.25,  non-bessemer,  $3.50.  The  difference 
in  price  between  the  Old  Range  and  the  Mesabi  ore  is  because  the 
latter  is  soft  and  contains  much  fine,  and  thus  is  less  acceptable 
for  blast-furnace  work.  Bessemer  ore  is  low  in  phosphorus,  not  to 
exceed  0.045%  phosphorus  to  55%  iron. 

Pig-iron,  Pittsburg  market  (Quotations  for  carload  lots). — Stand- 
ard bessemer,  $15 ;  malleable  bessemer,  $14.50 ;  basic,  $14.25 ;  No.  2 
foundry,  $14.50;  gray  forge,  $13.50.  Standard  bessemer  is  used  for 
making  steel  in  the  bessemer  converter,  malleable  bessemer  for  malle- 
able iron  castings,  basic  for  steel  suited  to  the  basic  open-hearth  furn- 


484  THE 

ace,  foundry  for  making  foundry  castings  suited  to  machining,  and 
gray  forge  for  wrought-iron.  In  the  Chicago  markets,  both  Southern 
and  Northern  pig-iron  are  quoted.  The  first,  from  the  great  iron 
center  at  Birmingham,  Alabama,  though  cheap,  is  high  in  phos- 
phorus. The  Northern  iron  from  nearby  points,  is  made  from  Lake 
Superior  ores.  Pig-iron,  cast  in  sand,  is  weighed  to  2260  Ib.  for  a 
long  ton,  the  20  Ib.  excess  being  an  allowance  for  the  sand  that 
sticks  to  the  pigs. 

Steel,  Pittsburg  market. — Bessemer  and  open-hearth  billets  are 
quoted  at  $25.  These  are  ingots  4  in.  square  by  6  ft.  long  that  are 
re-heated  and  rolled  into  the  required  merchant-steel  bars.  Mer- 
chant-steel bars  remain  at  1.40  to  1.60c.  per  pound.  This  is  equal 
to  $31.36  to  $35.84  per  long  ton.  Structural  steel  includes  angles, 
channels,  and  I-beams,  and  is  quoted  much  as  are  merchant-steel 
bars.  Steel  sheets  are  quoted  at  2.50c.  for  black  and  3.35c.  per 
pound  galvanized,  No.  28  gauge.  The  'black'  means  that  the  sheets 
are  rolled,  but  not  galvanized.  The  gauge  referred  to  is  the  No.  28 
wire-gauge,  and  is  the  basis  from  which  thickness  is  reckoned  accord- 
ing to  a  fixed  scale.  Steel  railroad  rails  are  held  at  $27.  Scrap  or 
old  steel  material,  held  at  $10  to  $12  per  ton,  is  added  to  the  charge 
in  basic  open-hearth  work. 

Silver. — At  New  York  the  quotations  are  on  silver  bars,  per  troy 
ounce  of  silver,  1000  fine.  It  takes  14.58  troy  ounces  to  make  1  Ib. 
avoirdupois.  London  prices  are  for  sterling  silver,  925  fine.  The 
value  of  the  pound  sterling  is  also  given,  so  that  with  the  London 
quotation,  we  may  compute  the  equivalent  price  in  cents  there.  Let 
us  say  that  sterling  exchange  is  $4.86,  and  that  silver  is  selling  at 
25l  per  sterling  ounce,  we  have  then : 

i^S^S"  ==  54-8c-  Per  ounce  of  fine  silver • 

Copper. — The  New  York  price  is  expressed  in  cents  per  pound, 
quotations  being  given  for  Lake  copper  cast  in  the  form  of  cakes 
for  rolling  into  sheets,  or  ingots  for  re-melting  to  make  castings, 
brass,  and  bronze,  or  wire-bars  for  drawing  into  wire.  A  sample- 
quotation  is  as  follows:  "The  market  closes  steady  at  13%c.  for 
Lake,  131/4@13%c.  for  electrolytic."  (Here  is  noticed  a  difference 
of  %  to  y±Q.  per  Ib.  in  favor  of  Lake  copper).  "Casting-copper  has 
averaged  13@13%c.  during  the  week."  Casting-copper  is  not  as 
pure  as  that  which  is  to  be  rolled  into  sheets  or  drawn  into  wire. 
Electrolytic  copper  is  made  by  re-melting  cathodes  (the  product  of 
electrolytic  refining)  into  ingots,  cakes,  or  wire-bars.  Cathodes  are 
held  at  %c.  less  than  electrolytic  copper,  the  difference  paying  for  the 


OF    THE    COMMON    METALS.  485 

re-melting.  In  London  copper  is  sold  by  the  long  ton  in  English 
money,  and  is  of  various  brands.  A  sample  set  of  prices  is  as  follows : 

English  tough  copper,  £63  10s. 

Best  selected,  £62  10s.  @  £63  10s. 

Standard,  £59  17s.  6d.  for  spot ;  £60  13s.  for  3  months. 

The  last  quotation  has  reference  to  whether  the  copper  is  for 
immediate  delivery,  or  whether  the  customer  will  take  it  at  the 
expiration  of  three  months,  in  which  time  the  reduction-works  will 
have  produced  it.  The  making  of  best-selected  copper  is  mentioned 
under  'the  refining  of  copper'.  Standard  copper,  formely  called 
g.  m.  b.  (good  merchantable  bars),  is  the  grade  upon  which  the 
others  depend.  Besides  these  brands  we  have : 

Strong  sheets  (rolled  copper) £72  10s. 

India  sheets  (a  rolled  brass) £68  10s. 

Yellow  metal  (a  grade  of  brass) 5%d.  per  Ib. 

It  is  the  business  of  dealers,  and  others  interested  in  copper,  to 
keep  statistics  of  the  supply  of  available  copper,  which  is  called  the 
'visible  supply'. 

When  the  visible  supply  is  small  the  price  naturally  rises,  and 
the  reverse  is  true  when  it  is  large.  Manufactured  copper  is  quoted 
at  19c.  per  Ib.  for  cold-rolled,  and  18c.  for  hot-rolled,  while  wire  has 
a  basis  price  of  15i4c.  in  carload  lots  at  the  mill. 

Tin. — Like  copper,  tin  is  quoted  for  immediate  or  for  future  de- 
livery at  a  specified  time.  A  sample  quotation  would  be  30%c.  for 
spot  and  29%  to  30c.  for  future  (three  months,  for  example).  The 
price  is  the  same  whether  the  metal  is  from  Burma,  near  the  Straits 
of  Malacca  (Straits  tin),  from  Bolivia,  from  Australia,  or  elsewhere. 
When  sold,  as  in  the  London  market,  we  quote  £134  for  spot,  and 
£130  15s.  for  three  months  per  long  ton  of  2240  pounds. 

Zinc  is  commercially  called  spelter.  Quotations  in  the  United 
States  are  given  in  cents  per  pound,  thus :  4.60c.,  St.  Louis ;  4.75  to 
4.80c.,  New  York.  The  London  market  is  quoted  at  £19  15s.  for  good 
ordinaries  (ordinary  brands)  and  £20  for  specials  (the  purer  zinc). 
St.  Louis  is  near  the  zinc-producing  district  of  Kansas,  Missouri,  and 
Illinois,  and  hence  has  a  lower  price  for  spelter  than  New  York. 

Prices  for  zinc  ore  are  given  per  short  ton  at  Joplin,  Missouri,  on 
a  basis  for  ore  assaying  60%  zinc,  and  the  price  varies  with  the  zinc 
content  above  or  below  this  figure.  It  once  was  the  custom  to  obtain 
the  basis  price  per  ton  in  dollars,  by  multiplying  the  price  of  spelter 
per  pound  in  cents  by  7.5,  but  this  is  only  an  approximation  the 


486  THE    METALLURGY 

buyer  often  being  willing  to  pay  a  high  price  to  secure  the  ore  when 
it  is  scarce.  With  zinc  quoted  at  5c.  per  pound,  the  price  for  zinc 
ore  would  be  $37.50. 

Antimony. — Sample  quotations  per  pound  are,  8l/±  to  8%c.  for 
Cookson's,  7%  to  8%c.  for  Hallett's,  and  ll/2  to  7%c.  for  ordinary 
brands.  The  brands  named  are  those  of  well  known  smelters  of 
antimony. 

Quicksilver. — The  New  York  price  is  $42.50  per  flask  (75  Ib.  net) 
for  large  lots.  San  Francisco  makes  nominal  prices  per  flask  at  $42 
for  domestic  orders,  and  $40  for  export.  The  London  price  is  £8  5s. 
per  flask.  Formerly  the  weight  of  a  flask  was  861/4  Ib.,  but  this  has 
been  decreased  to  75  pounds. 

Precious  metals. — Gold  is  sold  to  the  mints  at  an  unchanging 
price  of  $20.67  per  troy  ounce,  1000  fine.  From  this  the  mint  makes 
a  deduction  of  2c.  per  oz.  to  cover  the  cost  of  melting  and  assaying. 
Platinum  is  a  commercial  metal,  at  present  more  valuable  than  gold. 
We  quote  $22.50  for  hard  platinum,  $20  for  ordinary,  and  $16  for 
scrap.  Silver  has  already  been  discussed. 


INDEX 


Page. 

Accounting  department .......  478 

Acid  bessemer  blow 444 

Refractories 48 

Agitation    vats 211 

Alkaline  earths  in  slags 383 

Amalgam    safe 227 

Amalgamating  pan 222 

Amalgamation    118 

Breaking  ore  for 68 

Of  silver  ores 218 

American  ore  hearth 368 

Analyses  of  coke 385 

Of  copper  matte 315 

Of  dolomite 383 

Of  flue  dust 391 

Of  fuels.  .33,  34,  35,  36,  39,  43,  45 

Of  gold  ores 137 

Of  iron  slags 293 

Of  lead  ores 361,  362 

Of  limestone 383 

Of  matte Ill,  315 

Of   pig-iron 294 

Of   refractories..    50,  51,  54,  55 

Of  slags 382 

Of  silver  precipitate 259 

Anaconda,      Montana,      matte 

smelting   334 

Annealing  zinc  retorts 409 

Anode-mud 435 

Anodes,  sampling  of 65 

Anthracite    35 

Market    483 

Antimony  in  base  bullion 384 

Prices    486 

Argentite    217 

Arsenic  in  silver-lead  smelting  384 

Assayer,  duties  of 474 

Augustin  process 248 

Automatic    charging    of    blast- 
furnaces      276 

Feeders   126 

Sampling    59 

Azurite    298 

B. 

Bag  house 392 

Barrel  chlorination 143 

Base  bullion 390 

Bullion,  sampling  of .  . .  64 


Page. 

Metal  leaching. 255 

Metal  ores 18 

Bases  in  slags 382 

Basic  open-hearth  process 444 

Refractories 48 

Bedding  lead  ores 362,  369 

Beehive  ovens 39 

Belt  elevators 458 

Berthelot  law 22 

Bessemer  converter 442 

Iron   ores 18 

Process    441 

Best-selected  copper 332 

Betts    process 448 

Black  Pine,  Nevada,  treatment 

at    243 

Blaisdell  excavator 183 

Biake    crusher 69 

Blast  for  copper 303 

Roasting    112 

Blast-furnace,  breaking  ore  for  66 

For  silver-lead 371 

Plant  275 

Blende     399 

Blister    copper 332 

Blowing    engines 283,  285,  346 

Blowing-in 286,  377 

Bodie,  California,  treatment  at  209 

Bone-ash    48,  54 

Bornite    297 

Bosqui,     F.     L.,     acknowledg- 
ments to 13 

Boss  process 229. 

Bottoms,  treatment  of 333 

Boulder  county,  Colorado,  gold 

ores 137 

Breaking    ore 66 

Brick   kiln 49 

Machine 53 

Mold 52 

Briquetting    press 394 

British  thermal  unit,  definition  20 

Bromo-cyanogen  process 210 

Brown    agitator 265 

Brown-horseshoe   furnace 97 

Brown-O'Harra    furnace 97 

Bruckner    furnaces 97 

Butters   distributor 182 

Filter    J86 


488 


INDEX. 


Page. 

By-product  charcoal 37 

Ovens 39,  40 

C. 

Calamine    399 

Calciners    92 

Calcining    331 

Calculation  for  matte  charge. .  324 
Of  charge  in  pyrite  smelt- 
ing      322 

Of    charge    in    silver-lead 

smelting    385,  389 

Of  iron-furnace  charge...  291 

California  gold  ores 137 

Stamp-mills  133 

Callow  cones. , 178 

Calorie,  definition 20 

Calorimeters    22 

Cams    124 

Cananea    ores 137 

Canda    cam 124 

Canyon    City,    Colorado,    lime- 
stone      383 

Capacity  of  roasting  furnaces  111 

Of    stamps 133 

Carbonates  of  copper 298 

Carbon-brick    48,  54 

Cast-iron    refining 439 

Cast-steel  ladle 348 

Casting-lead 424 

Machine    287 

Pig-iron     287 

Centrifugal   pump 173 

Cerargyrite    217 

Chalcocite   297 

Chalcopyrite    297 

Charcoal 36 

In  silver-lead  smelting...  385 

Charge  floor  in  lead  smelting. .  378 

For  zinc  smelting 407 

Scoop   for  zinc 406 

Sheet  for  iron  furnace. . . .  292 
Sheet  for  matte   smelting  " 

315,  324 

Sheet  for  pyrite  smelting.  322 
Sheet  for  silver-lead  smelt- 
ing          387,  389 

Charging  cyanidation  vats 164 

Chemical    reactions    in    blast- 
furnace   .  288 


Page. 

Chamist,   duties   of 474 

Chemistry   of  blast-furnace...  381 

Of  cyanide  process 154 

Of   roasting 79 

Of  roasting  zinc  ores 400 

Chilean   mill 207 

Chimneys 28 

Chlorination    36,  135,  139 

Chloridizing  ores 18 

Roasting  of  silver  ores 235 

Chlorine  generator 140 

Chrome-iron   48,  51 

Chrysocolla    298 

Cinder  from  blast-furnace 286 

Classification     of     cyanidation 

methods    156 

Of  gold  ores 115 

Classifiers 177,  189,  201 

Clean-up  at  Maitland  mill 203 

In  Cyanidation 170 

In  silver  cyanidation 270 

Pan 128 

Coal 31,  34 

Market    483 

Coarse  crushing 69 

Coke    31,  38 

In  silver-lead  smelting.  . . .  384 

Market    483 

Coking  coal 34 

Combination  M.  &  M.  Co.,  treat- 
ment at 243 

Silver  mill 234 

Combustion    24 

Commercial  considerations. .  . .  467 

Composition  of  matte 315 

Comstock  ores,  treatment  of . . .  219 

Concentrate,  cyanidation  of...  209 

Definition  18 

Purchase  of 469 

Roasting  of 244 

Sampling  of 63 

Condensers  for  zinc  smelting.  .  408 

Coning 58 

Construction   of   reverberatory 

furnaces    94 

Work 457 

Converter,  bessemer 442 

Plant  for  copper 342 

Converting  copper  matte 336 


INDEX. 


489 


Page. 

Conveyor  belts 461 

Copper    295 

Converter   336 

In  cyanidation 208 

In  lead  smelting 384 

Ingots,  sampling  of 65 

Ores,  purchase  of     469,  471,  472 

Quotations    484 

Refining    430 

Sulphides,  pot- roasting  of.  113 

Sulphides,  roasting  of 84 

Cost  at  Lexington  mill 242 

Of  chlorination 142,  150 

Of  converting 349 

Of  cyanidation..      191,  193,  210 

Of  electrolytic  refining. . .  .  437 

Of  gold  ore  treatment 213 

Of  grinding 74 

Of  heap  roasting 90 

Of  labor 477,  480 

Of  lixiviation    260 

Of  mechanical  roasting...  Ill 

Of  refining  base  bullion. . .  430 

Of  reverberatory  furnaces.  93 
Of  roasting    in    reverbera- 

tories 96 

Of  roasting  stalls 90 

Of  Russell  process 262 

Of  sampling 63 

Of  silver  cyanidation 270 

Of  silver  milling 229 

Of  smelting  lead  ores 395 

Of  stamp  milling 134 

Of  zinc  smelting 410 

Cowper  stoves 281 

Cripple  Creek  ores 137 

Ores,  cyanidation  of 205 

Ores,  treatment  of 143 

Crushing 66 

For  cyanidation 151 

For  patio  process 246 

Lead  ores 364 

Cupola  furnace 26 

Cupelling    427 

Furnace 426 

Cuprite    298 

Current  density  in  refining. . . .  435 
Cusihuiriachic,    Mexico,    treat- 
ment at. .  262 


Page. 

Cyanide,  Colorado,  treatment  at  205 

Cyanidation  of  gold  ores..   135,  151 

Of  silver  ores 263 

D. 

Dale,  California, -process  at...  208 

Dead  fluxes 386 

Decantation     156,  180,  185 

Diehl    process 210 

Direct    process 333 

Discharging  cyanidation  vats.  165 

Discipline     475 

Distillation,  breaking  ore  for.  68 

Of  zinc 402 

Dolomite     48,  54,  383 

Double-discharge    mortars....  123 

Treatment  in   cyanidation 

153,  181 

Draft  in  chimneys 28 

Dressing    plates 127 

Drop  of  stamps 133 

Dry-crushing,  flow  sheet  for..  73 

Silver    mill 238 

Dry    ores 18 

Ores,   purchase  of 469 

Silver  milling 238 

Drying  cyanide   precipitate...  173 
Ducktown,    Tennessee,    smelt- 
ing   practice 321 

E. 

Edwards    furnace 97,  101 

Electric  trolley  slag  pots 326 

Electrolyte,  circulation  of 437 

Purifying    436 

Electrolytic  copper  refining.  . .  434 

Refining  of  lead 448 

El  Oro  cyanidation  method...  193 

Elevator  buckets 459 

Enargite   298 

Endless-chain  conveyor 462 

Endothermic    reactions 21 

English  cupelling  furnace....  426 

Equipment  of  plant 453 

Extraction    19 

And  mesh  in  cyanidation.  194 

By  Russell  process 262 

In  silver  milling 235 

Of  copper 299 


490 


INDEX. 


Page. 

Of  copper  from  copper  sul- 
phides      350 

Of    silver 218 

Exothermic  reactions 21 

F. 

Faber  du  Faur  retort 425 

Feeders     126 

Filter  presses 171 

Pressing    172,  185 

Pressing     in     cyanidation 

156,  188 

Fine  crushing 72 

Fineness  of  gold 486 

Firebrick    51 

Fire-clay    48,  51 

Flash   roasting 137 

Flow-sheet  in  Maitland  mill. . .  200 

Of  Augustin  process 249 

Of  barrell  chlorination. . . .  151 

Of  dry-crushing 73 

Of  lead  refining 417 

Of  Russell  process 261 

Of  Ziervogel   process 250 

Flue  dust 391 

Gases,  temperatures 28 

Fluorspar  in  slags 383 

Fore-hearth    309 

Foreman,  qualities  of 474 

Fractional  selection  of  samples  59 

Free-milling  ores 18 

Fuel 31,  43 

In  silver-lead  smelting...  384 

Market 482 

Furnace,  cupola 26 

English    cupelling 426 

Reverberatory  matting....  327 

Wind   26 

Zinc    404 

Fuse    box 93 

Fusion  of  copper  ore 331 

G. 

Galena 361 

Roasting  of 112 

Gangue,  definition 18 

Ganister 48,  49 

Gates  tube-mill 196 

Genesis  of  fuels 32 

Globe  bag  house 393 

Plant  .  370 


Page. 

Gold  bars,  sampling  of 64 

Metallurgy  of 115 

Native   115 

Silver,  parting 439 

Solution  tank 168 

Tellurides 115 

Golden  Cycle  mill 207 

Grabs 464 

Grab   samples 57 

Grading  ore 19 

Pig-iron 294 

Graphite   35,  48,  50 

Grinding    66,  74 

H. 

Hand  reverberatories 92 

Sampling  57 

Handling  materials 457 

Harrisburg,  Arizona,  treatment 

at    210 

Heap-roasting,  breaking  ore  for 

67,  84 

Heat  evolved  in  reactions 21 

Of  formation  of  compounds  23 

Reactions  in  roasting 81 

Hematite 273 

Henderson  process 356 

Hersam,    E.    A.,    acknowledg- 
ments to   14 

Heyl    and    Patterson    casting- 
machine    287 

Hoffman  by-product  oven 41 

Hoists    464 

Holthoff    furnace 97 

Homestake    cyanidation    meth- 
ods       188 

Hot-blast     stoves 282 

Howard   mixer 421 

Howard     press 422 

Hunt  ammonium  process 208 

Hunt     and     Douglas     process 

350,  353 

Huntington-Heberlein  process.  112 

Hydraulic   accumulator 347 

Hydro-metallurgy   of   copper..  349 

Of     gold 134 

Of    silver 248 

Hyposulphite  treatment  of  sil- 
ver  ores 253 


INDEX. 


491 


Page. 


I. 


Impurities  in  lead  bullion....  416 

Industrial    railways 464 

Ingots,   sampling   of 64 

Installation   of   plants 455 

Iron    in   slags 382 

Ore,    market 483 

Ores     273 

Ores,   purchase   of 460 

J. 

Jeffrey    elevators 459 

K. 

Kiln,    brick 49 

For  making  charcoal 37 

Krupp    ball-mills 137 

L. 

Labor  costs 480 

Lake  copper,  refining 433 

Lathe  for  cutting  zinc 169 

teaching  base  metals 255 

Grinding  ore  for 68 

In    cyanidation 159,  165 

In   Ziervogel   process 252 

Ores   18 

Silver    ores 257 

Lead  bullion,   refining 415 

Copper  matte 395 

Electrolytic    refining 448 

Ores    361 

Ores,  purchase  of 469 

Leadville    ores 361 

Lexington  mine,  treatment  at.  241 

Lignite    33 

Limestone    383 

Lining   converters 341 

Lixiviation  of  silver  ore 253 

Location   of   works 453 

Loomis-Pettibone  gas-producer.  45 

Loss  in  converting  copper....  349 

In    distilling    zinc 409 

Losses   in   roasting Ill 

Lump,  ore,  roasting  of 84,  89 

M. 

Machine     sampling 59 

Magnesite    48,  54 

Magnetite     273 


Page. 

Mahler     calorimeter 22 

Maitland  cyanidation  mill.....  199 

Malachite     298 

Manganese  in  slags 382 

Market    lead 420 

Matte     314 

Analyses  of Ill 

Composition    of 315 

Concentration      324 

Heap-roasting   of 87 

Lead    copper 395 

Smelting    302,  304 

Smelting  at  Anaconda.  . . .  334 

"Smelting,   charge  sheet...  316 
Treatment     by     Augustin 

process    249 

Treatment     by     Ziervogel  . 

process 250 

Matting    blast-furnace 303 

McArtnur-Porrest   process     135,  152 

McDougall  furnace    ...     91,  97,  108 

Mechanical   roasters 97 

Melaconite    298 

Melting  furnace  for  silver.  .  .  .  228 

In   lead 378 

Lake    copper 433 

Lead    ores 364 

Silver     residues 245 

Mercur,       Utah,       cyanidation 

method    156 

Mercury,  see  quicksilver 

Merton    furnace 97 

Mesh  and  extraction  in  cyan- 
idation       194 

Metal   market 482 

Metallic  Extraction  Co.,  treat- 
ment      205 

Metallics,    sampling 63 

Metallurgical  treatment  of  ores  19 

Metallurgy  of  zinc 400 

Metals,  sampling  of 64 

Mill-sites    455 

Missouri  dolomite 383 

Moisture    sample 57 

Molding  market  lead 424 

Monadnock     mill 207 

Mount  Morgan  ores,  treatment 

of    137 

Muffle  roasting  furnace...  358 


492 


INDEX. 


Page. 


N. 


Native    copper 298 

Silver    • 217 

Natural   gas 31,  36 

Solid    fuels 33 

Neill    process 355 

New  York  metal  market 483 

Nitric-acid   parting 439 

Non-bessemer  iron  ores 18 

O. 

Open-hearth  process 444 

Operating   department 473 

Operation  of  bessemer  conver- 
ter      443 

Of  blast-furnace  283 

Of  converters    338 

Of  copper  furnace 86 

Of  lead  smelting 368 

Of  reverberatories 94 

Of  stamp-battery    125 

Ore-buying  schedules 469 

For  matte  furnace 312 

Hearth     368 

Ores,   classification   of -17 

Definition  of 17 

Mixed    17 

Of  copper   297 

Of  gold    117 

Of  iron    273 

Of  lead    361 

Of  silver 217 

Of  zinc 399 

Simple    17 

Organization    of   metallurgical 

company     467 

Oxidation  of  lead  ores 366 

Oxides  of  copper 298 

Oxidizing  roasting 79 

P. 

Parkes  process 420 

Parting   gold-silver 439 

Patera   process 253 

Patio    process 246 

Pattison    process 419 

Pearce-turret    furnace     97,  104,  137 

Peat  33 

Petroleum   as  fuel 31,  35 


Page. 

Pig-iron,    market 483 

Sampling    of 66 

Smelting     274 

Pipe    sampler $6 

Pittsburg    markets 483 

Plant  and  equipment 453 

Plattner  process 135,  139 

Plumbago   50 

Polybasite    217 

Pot-roasting 112 

Pound-calorie,  definition 20 

Precipitation  after  chlorination  148 

After    lixiviation 258 

At  Maitland  mill 203 

In    cyanidation 169,191 

Preparation  of  ores 19 

Prices  of  pig-iron 294 

Producer    gas 4& 

Profits   481 

Properties  of  zinc 399 

Puddling     441 

Pulp,  sampling  of 63 

Pulverized  ore,  roasting  of...  91 

Punch    sampling 65 

Purchase  of  ores 468 

Push   conveyor 463 

Pyrite  matte  smelting 318 

Roasting     79 

Smelting  in  two  stages.  .-.  321 

Q. 

Quality    of   labor 477 

Quartering 58 

Quicksilver  fed  to  battery 126 

In    silver    milling 244 

Prices    486 

Trap    127 

Quotations    of    metals 482 

R. 

Rabbles  in  Edwards  furnace..  107 
Rand     cyanide     practice     153, 

156,  180 

Raymond   furnace 97 

Reactions  in  basic  open-hearth 

process    447 

In  bessemer    converter...  444 
In  blast-furnace    ....     283,288 

In  bromo-cyanidation     . . .  213 

In  chlorination    141 


493 


353 

257 
380 
247 
320 


Page. 

In  copper  converter 339 

In  cyanidation   154 

In  Hunt  and  Douglas  pro- 
cess     

In  leaching    silver    ores.. 

In  lead    smelting 366, 

In  patio    process 

In  pyrite    smelting 

In  reverberatory     matte 

smelting    330 

In  Rio  Tinto  process...  351 

In  roasting    8<< 

In  roasting  concentrate...  246 
In  roasting      copper      sul- 
phides      84 

In  roasting  silver  ores. . . .  236 

In  roasting  zinc  qres 400 

In  Russell  process   262 

In  silver  milling 224 

In  Ziervogel  process 251 

Receiving  lead  ores 362 

Ore   56 

Reduction  of  lead  ores 366 

Reese   river  process 238 

Refining    413 

Refractory   materials 47 

Re-lining  converters 341 

Replacing  zinc  retorts 407 

Retorts  for   refining 425 

For  zinc  smelting 405,  408 

Retorting    amalgam.     129,  22S,  245 

Reverberatory  lead  smelting..  365 

Matte    smelting. 326 

Roasting    furnalces 92 

Smelting,  breaking  ore  for  67 

Revolving    screen 74 

Rio  Tinto  process 351 


Page. 


S. 


Re-pressing    machine. 
Roasting     

Blende     

Breaking  ore  for. 


50 

....  77 

400 

67 

Concentrate    245 

For  Ziervogel  process...,  251 

Furnaces  for  zinc  ores. . . .  401 

Gold  ores 137 

Silver    ores 135 

Rolls    72 

Russell    process 260 


Sadtler    process 411 

Sampling    55 

Base   bullion 390 

Lead    ores 362 

Works   61 

Sand    48 

In  silver  cyanidation 267 

Treatment      at      Maitland 

mill    201 

Schedule  of  copper  ores 472 

Scoop  for  zinc  charging 406 

Screen   for  stamp-mills 124 

Revolving     74 

Tests    69 

Screw-conveyors     4C2 

Segregation    19 

Selling  department 482 

Settler  for  silver  milling 225 

Settling    tanks 177 

^Siderite     273 

Silica    brick 48,  49 

Silicates  of  copper 298 

Silver    215 

Bars,  sampling  of 64 

Lead  ores,  purchase  of...  469 

Lead    smelting 369 

Milling    218 

Quotations    484 

Single-discharge    mortars 122 

Site  for  works 454 

Skilled    labor 475 

Skimmer    423 

Slag    analysis 293,315,382 

Floor  in  lead  smelting...  379 

From  blast-furnace 286 

From  copper  furnace 325 

In  silver-lead  smelting...  381 

Pot    310,  326 

Slime  in  silver  cyanidation...  268 

Plants    192 

Treatment      at      Maitland 

mill    202 

Sliming,   grinding  for 74 

Smelter  plant  for  copper 342 

Smelting    copper    ores 299 

For   pig-iron 274 

Gold  ores 213 

Lead    ores..  365 


494 


INDEX. 


Page. 

Ores   18 

Silver-lead   ores 369 

Zinc   ores 402 

Softening  furnace 418 

Lead  bullion 41b 

Sombrerete,  Mexico,  treatment 

at    262 

Specific  heat 30 

Speiss    384 

Spelter   refining 437 

Sphalerite    399 

Spitzkasten    177 

Split    shovel 59 

Stacks   28 

Stall-roasting,      breaking      ore 

for     67,  84 

Stamp-mills   118 

Men    for 478 

Stamper    406 

Standard    plant,    Bodie,    Cali- 
fornia      209 

Starting   blast-furnace 312 

Steel  cyanidation  vats 163 

Making    441 

Market    484 

Stephanite    217 

Stetefeldt  furnace  97 

Stirring    paddle 430 

Suction    filtration 185 

Sulphatizing  roasting  of  matte  251 

Sulphide  ores 18 

Roasting  79 

Sulphuric-acid  parting 439 

Supplies,   purchase   of 472 


T. 

Tailing,  leaching  of 

Purchase  of 

Sampling  of 

Taylor  gas-prouucer 

Tellurides  of  gold 

Temperature  of  combustion. . 
Testing  current  for  refining. 

Tetrahedrite  

Thermo-chemistry  

Time  in  cyanidation 

Tin  quotations 

Tram  tracks 

Traveling  cranes 


159 

469 

63 

44 
115 

29 
437 
298 

20 
167 
185 
464 
464 


Page. 
Treadwell  stamp-mill,  costs  at     480 

Trench  sampling 58 

Tripper   461 

Trommel     ?4f 

Tube-mills     75,  196 

Tuyere    311 


U. 


Unskilled    labor. 


475 


V. 


Vacuum-filtration    in    cyanida- 
tion        156,  185 

Vats   for  cyanidation 160 

For  hyposulphite  leaching  256 

Vezin    sampler 60 

Vibrating  trough  conveyor.  .  .  .  463 

W. 

Washoe  process 219 

Water-jackets     307,  375 

Weighing   ore 56 

Weight  of  stamps   133 

Welsh    process 330 

Western    Australia    ore    treat- 
ment         210 

Wet  pan 345 

Silver  mill   220 

Wethey    furnace 97,100 

White    briquetting   press.     206,  394 
White-Howell  roaster.     97,  137,  239 

White   metal 332 

Wind  furnace 26 

Wood   32 

Wooden  cyanidation  vats 161 

Wrought-iron    manufacture. . .     440 

Z. 

Zince  boxes 167 

Furnace    404 

In    slags 383 

Lathe     169 

Metallurgy  of 400 

Ores   399 

Quotations    •  485 

Zone  of  fusion  in  blast-furnace  284 
Of    preparation    in    blast- 
furnace      284 

Of  reduction   in  blast-fur- 
nace  .  ...  284 


YC  634 


^ 


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